Search Documents

By H Holmstr÷m, A Widerlund, M Lindvall

At the Kristineberg mine in northern Sweden sulfide-rich, poorly buffered tailings have been deposited in five impoundments located along a valley. The tailings area has been remediated by applying a combination of till cover and flooding. A new method that has been applied to parts of the tailings area is to raise the groundwater table to saturate the tailings. By increasing pH by liming, thereby reducing metal mobility, a water-covered downstream impoundment is functioning as a trap for metals released from the till-covered impoundments upstream. In this paper the efficiency of the system approach used at Kristineberg is evaluated. Progressive changes of water quality from groundwater in the upstream till-covered impoundments via surface streams within the pond system to the water in the flooded downstream pond and its outlet are described. The results show that the system approach used in the design of the remediation programme of the tailings works as planned. The sulfide oxidation rate has decreased in the till-covered impoundments, and, as predicted, there is an out-flush of old weathering products which will continue for at least five years before equilibrium is reached. The groundwater in the till-covered impoundments has high concentrations of Fe, S, Mg and Zn, but the concentrations of metals such as Cu, Cd and Pb are much lower than before the till cover was applied. The limed pond is an efficient metal trap, with concentrations as high as 2.76 per cent Zn and 0.28 per cent Cu in the calcite-gypsum sludge developed on the tailings. The outlet from the flooded pond, and thus from the whole mining area, releases small amounts of metals to the recipient in comparison to the amounts that are in movement within the mining area. During the period May to November, 13 kg dissolved Zn and 111 kg particulate Zn was released to the recipient, whereas a total of 27 000 kg was transported by surface water within the impoundment system during the same period. The results from Kristineberg show that when planning remediation of complex mining areas, it is useful to adopt a system approach in which different methods are used on different parts of the area along the flow path of the drainage water, whereas flooding of the downstream pond is the key to the performance of the entire system.

Jan 1, 2003

By K S. Smith, B P. Jackson, P J. Lamothe

The upper Arkansas River south of Leadville, Colorado, USA, contains deposits of fluvial tailings from historical mining operations in the Leadville area. These deposits are possible non-point sources of acid and metal contamination to surface- and ground-water systems. We used stable-metal isotopes to help ascertain metal retention and release mechanisms that influence metal transport from the deposits to shallow ground-water systems. To accomplish this, we excavated an intact core from a small fluvial tailings deposit and performed laboratory column experiments to examine the amount of metals leaching through the core. Deionised water was continuously applied to the top of the core, and effluent was collected from the bottom of the core for a period of about two months. Three saturation conditions were used, including unsaturated, partially saturated, and fully saturated. Reducing conditions developed upon partial and complete saturation of the core. During the unsaturated leaching phase, core effluents were acidic (pH 2.8 - 3.5) and contained elevated metal concentrations. During a portion of the unsaturated leaching phase, stable-metal isotope spikes and conservative tracers were simultaneously applied to the top of the core. The isotopes provide a means to delineate metal behaviour within the core. During the unsaturated leaching phase, the isotope spikes did not exhibit breakthrough with simultaneously added conservative tracers. However, some of the stable-metal isotope spikes did break through when reducing conditions occurred in the core (during the partial- and completesaturation conditions). In this paper, we use Cd as an example of the utility of stable-metal isotopes in geochemical studies. In the case of Cd, after the core was partially saturated, the 111Cd spike was released as evidenced by a change in the Cd isotope ratios in the effluent. This release continued during the fully saturated leaching phase, however, the total Cd concentration did not increase. These results suggest that the 111Cd spike was retained inside the core during the unsaturated leaching phase, and only partially released as reducing conditions developed. Results from this core-leaching experiment indicate there is a large reservoir of water-soluble material within the fluvial tailings deposit, which yields elevated metal concentrations and high acidity, and which may degrade adjacent ground- and surface-water quality. Use of stable metal isotopes in this study facilitated the determination of different metal-retention processes, metal-release processes, and metal sources in the fluvial tailings deposit in response to changing geochemical conditions.

Jan 1, 2003

Bacterial leaching of sulfide minerals are fairly well established in heap, dump and in situ leaching and the primary bacterial treatment of refractory gold concentrate using CSTR have gained industrial prominence in recent years. Three cultures, a mesophile T ferrooxidans (DSM583), a moderate thermophile Sulfobacillus acidophilus (strain YTF1) and a extreme thermophile (Sulfolobus strain BC65) were used for the biooxidation (pre-treatment) of the Porgera gold-bearing pyrite concentrate under optimum O2 and CO2 enriched air (30 per cent O2 and ten per cent CO2) based on volume ratios. The positive effect of using oxygen and carbon dioxide enriched air for the biooxidation of gold-bearing pyrite mineral was clearly demonstrated where an increase in Fe release rate by a factor of 2.8, 1.8 and 1.5 for DSM583, YTF1 and Sulfolobus respectively were observed compared with the shake flask tests. The lag phase observed with the other test conditions such as in shake flask or with air decreased by about 80 per cent for all three cultures. The average observed dissolved oxygen concentrations in the pulp for these tests were 8.7, 5.8 and 5.0 mg/L for DSM583, YTF1 and Sulfolobus respectively, indicating that O2 was not limiting in these tests. The pulp redox potential was observed to increase with time for all tests coincident with the leaching rate and reaching a maximum of 663 mV, however the maximum value differs for each strain, with DSM583 exhibiting the highest followed by YTF1 and Sulfolobus. Biooxidation of pyrite under optimum gas enrichment at different pulp densities showed near total Iron extraction occurred for pulp densities of three, ten and 20 per cent (w/v). At high pulp densities, more so at 40 per cent than 30 per cent (w/v), the leaching rate decreased rapidly for tests with YTF1 and Sulfolobus. Monitoring of dissolved O2 shows that O2 demand increased with increasing pulp density and the average dissolved O2 concentration observed for YTF1 and Sulfolobus at 40 per cent solids concentration was 1.0 mg/L and below, suggesting limiting oxygen and carbon dioxide conditions for these two cultures. The decrease in leaching rate observed at high pulp densities (30 and 40 per cent) under optimum gas enrichment was accompanied by low metal dissolution and low redox potential values which was attributed to shear stress due to attrition of cells by high solids concentration, toxicity of metal ions especially Ag and Fe and the formation of precipitates and jarosites as confirmed from XRD analysis of leach residues hence impeding the leaching process. Limiting conditions due to high pulp density and viscosity were only observed for YTF1 and Sulfolobus. The subsequent cyanide leaching for precious metals from the pre-treated pyrite concentrate in all cases shows the extent of gold and silver recovery increased with the degree of pyrite oxidation and Au and Ag recoveries in excess of 80 per cent were achieved compared with only 18.1 per cent and 35.9 per cent for gold and silver respectively when treated by direct cyanide leaching on the original, untreated concentrate.

Jan 1, 2004

By Shinnosuke U, Zhang HE, Qian X

In this paper it is recommended that pyrite is to be separated as froth from arsenopyrite by selective depression of arsenopyrite in strong alkaline medium (1), by oxidation with chlorinated lime (II), by co-use of chlorinated lime with sodium carbonate (III), and by modification with `L' reagent (know-how) without filtering to get rid of residual copper ions after Cu activation (IV). The corresponding principles of these (new) effective methods were briefly discussed. The artificial mixture (1:1) of pyrite and arsenopyrite could be successfully separated, and concentrates of pyrite containing 2.12 per cent (I), 7.00 per cent (II), 2.60 per cent (III) and 3.16 per cent (IV) arsenic with recovery of 90.9 per cent (1), 89.5 per cent (II), 89.0 per cent (III) and 81.4 per cent (IV) respectively can be obtained by only one flotation. Some effective precautions were suggested to improve the separation of arsenopyrite and pyrite (or other sulphides) by flotation.

Jan 1, 1993

By J Nixon

I have to say that my story is not too much different to many others who grew up and worked in the great mining fields like Broken Hill, Mt Isa, Rosebery, Kalgoorlie or Kambalda and any one of them could have stood up here and spun you a good yarn.What does my paper offer you in this volume? By sharing my experiences I want to encourage you to consider your own experience with leadership, taking on change, learning from and contributing to the development of others.I have been very fortunate to be part of many new innovations into underground mining over the past 40 or so years; most of it with Conzinc Rio Tinto Australia (CRA)/Rio Tinto operations.•From hand-held drills and a homemade single boom Jumbo mounted on an old Airtrak, to the sophisticated twin boom jumbos with high speed drifters all controlled by an on-board computer;•from a Scraper winch and a Wagner 2D loader to fully automated electric LHDs that navigate to the tipple, empty the bucket and return to where they come from;•from hand-held cut-and-fill and timber stoping in Broken Hill to the bulk mining methods of a block cave where at Northparkes we mined more ore in a month than we mined in the total five years of operations at Currawang;•from hand-held scaling and bolting with no skin protection to installing ground support with Jumbos and sophisticated Cable Bolters and the introduction of in cycle shotcrete;•from hand-held ladder rising and shaft sinking to Box Hole rigs, Raise Bore Machines and Blind Hole shaft sinking; and•from Gelignite based AN60 stick explosives, ANFO, fuse firing and half second detonators to Powergel, Emulsions and electronic computer programmed detonators that can somehow be initiated by ‘Blast Ped’ travelling through the rock.

Mar 21, 2011

Botanical exploration has identified many plant species that naturally accumulate very high concentrations of elements, particularly heavy metals. The enormous potential applications of some of these species to remove metals from soil (phytoextraction) are slowly being realised. The discovery of hyperaccumulation has resulted in considerable research into applications for removing contaminants (phytoremediation) or more importantly, in the context of this discussion, for commercially æfarmingÆ metals (phytomining) from metal-bearing soils. Currently published phytoextraction models estimate Ni yields using Berkheya coddii (a South African nickelophyte) are approximately $US 750 ha-1, excluding the cost of processing. Research indicates that B. coddii will also accumulate significant amounts of Co, resulting in a considerable increase in the value of the crop. Modification of the substrate by acidification, fertilisation and/or chelation indicate fruitful avenues for improving this yield by increasing the extractability of the target element, elevating biomass production rates and/or rendering other contained heavy metals extractable. Plant uptake of more insoluble elements, such as Au, may be induced in many plant species using chemical compounds. These chelates applied to the soil, form soluble complexes with poorly soluble metals forcing them into the soil solution. Plants may then accumulate these metal complexes through transpiration. Research has achieved foliar Au concentrations of 78 mg kg-1 in Iberis intermedia, 56 mg kg-1 in Brassica juncea and 9 mg kg-1 in B. coddii (on a dry weight basis). Therefore in theory B. coddii could phytomine a soil containing modest concentrations of Ni, Co and Au to yield 116 kg Ni, 36 kg Co and over 6 oz Au ha-1 at a current value of $US 4000. Root crops have also shown promise in chelate assisted Au accumulation (eg 6.8 mg kg-1 in Beta vulgaris and 118.5 mg kg-1 in Raphanus sativus). Phytoextraction, in its varied applications, is rapidly coming of age as a technology having considerable potential for application in environmental science and the minerals industry. Phytotechnology employs unique plant biochemical behaviours in the rhizosphere, driven by the sunÆs energy, to seek out and concentrate low levels of heavy metals at very low cost.

Jan 1, 2002

By Siddall B

A brief review of the capabilities and practical limitations of blast simulation packages has been carried out. It is concluded that the simulators accurately reflect size distributions in the mid-ranges of 0.1 to 1.0 in, but are less reliable

Jan 1, 1998

By Revy T

A testwork program aimed to improve the assaying of underground face samples established that for an Accelerated Cyanidation Extraction using the LeachWELLTM 60X catalyst, followed by an Aqua Regia finish (to quantify non-cyanide-extractable gold): ò Assay results (including the aqua regia finish) from a full set of samples from one underground face can be returned within four hours. ò The reagent dose to make the method simple and 'bullet-proof' is 8 gm LeachWELLTM 60X and 20 gm NaCN per 200 gm sample (30 per cent solids). Pulverising is to 90 per cent passing 75 micron (compared to the standard of 90 per cent passing 150 micron), and cyanidation is completed within one hour - however both might be further optimised. ò The LeachWELLTM/Aqua Regia results demonstrate that Fire Assays suffer from a significant nugget effect which makes the Fire Assay data relatively unreliable. ò Original costs, specific to Bannockburn (viz including an Aqua Regia finish), handling approximately 20 000 to 25 000 per annum is estimated to be $5.40, ie about one-third less than fire assaying. Optimising is further reducing these costs. It is estimated that the LeachWELLT"'/Aqua Regia assay procedure has reduced the probable error limits of any given assay of this ore from the range 0.74x to 1.36x for Fire Assay, to an appreciably smaller range of 0.90x to 1.11x. Thus for a cut-off grade of 1.60 g Au/t, samples with Fire Assays ranging up to 2.18 g Au/t have 25 per cent chance of being below cut-off, while samples with Fire Assays as low as 1.18 g Au/t have an equal chance of actually exceeding cut-off grade. This 1 g Au/t range in uncertainty is generally unacceptable for many practicalgrade/tonnage relationships. By contrast, using LeachWELLTM + Aqua Regia assays, the estimated corresponding limits are 1.44 g Au/t and 1.78 g Auk respectively - a far more tolerable 0.34 g Au/t range of uncertainty. Subsequent data has demonstrated that sufficient accuracy for grade control is achieved by Accelerated Cyanidation alone, further halving the cost of assays.

Jan 1, 1994

-in 1961,.West Australian Petroleum's Eneabba. No. 1 bore encountered twenty-two' seams of coal totalling forty feet in thickn' ss between the..'depths of 6373 feet and 64,/,1 feet. 0n'ana:lysis; this coal had a. similar calorific value (8000 - 9000 B.

Jan 1, 1964

By Ricketts N. J, Smith G. D J

A hydrometallurgical process was developed for recovery of copper from copper-lead matte formed as an intermediate product at the BHAS lead smelter at Port Pirie. A plant utilizing this process was commissioned in 1984. The leach step involves oxidation of the copper-lead matte in an oxygenated acidic chloride-sulphate leachant. The results of laboratory scale testwork from which the chemistry of leaching was deduced, are presented in this paper. The rate of recovery of copper to solution from copper-lead matte depends largely_on the availability of oxygen to the sulphide surface. The optimum leach temperature is approximately 85¦C. Free acid must be supplied according to the stoichiometry of the leach reaction: Cu1.91 S + PbS + 2.91 H2SO4 + 1.455 02 = 1.91 Cu2+ + 1.91 SOj- + PbSO4(s) + 2.91 H2O + S¦ The grind size of the matte is of minor significance. Chloride concentration in leachant should be at or above a minimum of 10.0 g/l. Chloride acts to complex cuprous ions formed during leaching.

Jan 1, 1987

The effects of changes in sintering bed height, coke fuel, suction and basicity on sinter productivity, physical strength and mineralogy were determined. It was found that, for a given iron ore, there is a narrow -coke range which maximises sinter strength, irrespective of bed height (in the range 300 to 500 mm) or basicity (in the range 1.2 to 2.2). Increases in bed height, at a constant coke level, improved the physical strength of the sinter. Productivity increases were obtained by increasing the sintering suction but, at coke levels other than the optimum, these *ere somewhat offset by the decrease in sinter physical strength. At basicities higher than 1.2 (i.e. 1.6, 1.8 and 2.2) there was a narrowing of'the size range of product sinter, with a decrease in the percentage of top size material, and a decrease in the amount of fines produced. Changes in coke, bed height and basicity exert major effects on sinter mineralogy, while the Nagoya index was mainly influenced by cote and bed height Changes.

Jan 1, 1975

By K L. Brown

The concentrations of trace metals in deep hydrothermal solutions (~1 km depth, 200 to >300¦C) from modern hydrothermal systems in the Taupo Volcanic Zone, New Zealand, and Ladolam, Lihir Island, Papua New Guinea were obtained using a purpose-built downhole fluid sampler and ICP-MS analyses. The results range widely: gold (<0.1 - 23 ppb); silver (<3 - 2400 ppb); arsenic (100 - 18 000 ppm); antimony (2 - 1200 ppm); mercury (<1 - 78 ppb) and provide the basis for quantifying the flux and fate of metals as well as assessing their sources (Simmons and Brown, 2006, 2007). An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation.

Jan 1, 2008

By R Butler

Contract m1n1ng in New Zealand developed in the coal industry and has become an element in the success of recently opened hard rock gold mines. At Martha Hill, the mining contractor is undertaking a wide range of activities from open pit mining, crushing, conveying, waste embankment construction and drainage.

Jan 1, 1991

By Q Xin, X Shengming, C RuÆan, Z Zhigeng a

The optimum defluorinating conditions for Sichuan bastnasite crude ore were determined by orthogonal experiments. Over 95 per cent fluorine in the crude ore was stripped by roasting 30 minutes of the crude ore with Na2CO3/ore =0.25 at 650oC. The stripped fluorine crude ore was further treated by NH4 Cl to recover rare earth. Above 95 per cent rare earth can be leached by water after the stripped fluorine crude ore mixed NH4Cl(NH4Cl/ore =1:1) roasted 1 h at 325oC. At the same time, the rare earth leaching solution contains little Fe, Si and Al. Therefore it is propitious to further purify the solution and recover rare earth.

Jan 1, 2000

By P K. Seccombe

The McKinnons gold deposit is hosted by weakly deformed, Palaeozoic metasedimentary rocks of the Cobar Basin in the western Lachlan Fold Belt of New South Wales. Mineralisation comprised a largely supergene resource of 1.7 Mt of ore, grading 1.7 g/t gold, established over structurally controlled, primary gold and base-metal mineralisation in underlying, carbonate-rich meta-siltstones and sandstones of Early Devonian age. Mineralisation was associated with at least two major fluid cycles. The first cycle was characterised by four stages of vein formation. Early quartz and carbonate veins (Stages 1 and 2) and quartz-pyrite breccias (Stage-3) occurred prior to, or early in, the main regional deformation (D1). Continued D1 deformation produced syntectonic, carbonate-base-metal veins (Stage-4), primarily in dilatant zones adjacent to fold axes. The second cycle of mineralisation was an epithermal-style event, driven by a shallow, syn-D2 or post-tectonic intrusion, defined by aeromagnetic imaging. The intrusion may have exploited a jog, controlled along the basin margin by basement faults. Mineralisation produced gold-pyrite veins, associated with pervasive quartz-sericite-adularia and chlorite-sericite alteration. Highest primary gold grades are confined to quartz-adularia boiling zones. The McKinnons deposit offers an important example of mineralisation formed by multiple processes. Early-stage, base-metal mineralisation at McKinnons reflects structurally focussed, basin dewatering and metamorphic processes, typical of other deposits in the Cobar Basin. By contrast, late-stage gold mineralisation at McKinnons is magmatically driven and of epithermal style. Recognition of overprinting distinctive geological events is critical, not only in modelling ore formation, but in developing a versatile exploration concept.

Jan 1, 1999

During mapping in the Bijerkerno area (Fig. 1), well-preserved sedimentary structures were found at three separate localities. The structures are of four varieties, which are 1. small-scale cross-bedding (Fig. 2),2. convolute bedding (Fig. 3),3. minor scour and fill (Fig. 4), and4. flame structures.Ripple-marking is also doubtfully present.These structures are best preserved in what was...

Jan 1, 1971

By Segnit ER, Jensen G

The decoppering of soft and hard leads was investigated by stirring in an open pot, using sulfur or sulfur plus pyrites, or sulfur plus pyrrhotite, as reagents. Lead samples taken during the process were analysed for copper. Dross samples were examined by optical micro- scope, SEM and XRD. Soft lead containing silver can be decoppered satisfactorily by stirring with 0.1% S at low temperatures - close to 330¦C. Soft lead, whether containing silver or not, can be decoppered satisfactorily at higher temperatures by stirring with 0.1% S plus 0.4% pyrites. Then CuFeS2, Cu5FeS4 and PbS are formed, instead of Cu2S, CuS and PbS when using S alone. Hard lead containing tin can be decoppered satisfactorily by stirring with 0.1% S at temperatures from 320-400¦C. When stirring with 0.1% S plus 0.4% pyrites, reversion of copper from dross to lead on prolonged stirring does not occur. Hard lead without tin can be decoppered with S plus pyrites better than with S alone, but the presence of antimony hinders the decoppering, which is less effective on lead with 5.5% Sb than with 2.3% Sb.

Jan 1, 1980

By Quan P. M

The great advances made in the application of solvent extraction to the recovery of copper from low grade ore sources over the past 15 to 20 years are well known. Solvent extraction is an essential step in the recovery and purification of copper from dilute sulphate leach solutions by which some one million tonnes of copper are now electrowon annually. A reagent is also available for the selective recovery of copper from chloride leach solutions and this may hold the key to the processing of complex sulphide ore concentrates. By comparison, the use of solvent extraction in the recovery and processing of zinc has as yet found little favour and is confined to only a few specific and comparatively small applications. This is despite the ready availability of a large number of low grade zinc sources and zinc-bearing wastes and residues. It is suggested that this difference is due in large part to the absence of suitable reagents having the particular features required to enable them to fit into a solvent extraction - electrowinning based process for zinc. Ideally, for the recovery of zinc from dilute sulphate solution, an extractant should operate on a proton exchange principle and be strong for zinc in the same way that o-hydroxyaryloxime reagents are strong towards copper. It should load efficiently with zinc at the pH of typical sulphate feed streams without the need for neutralisation and yet be stripped efficiently by spent zinc electrolyte. Another factor of major importance for many applications, however, is that it should be highly selective for zinc over ferric iron. These features have now been realised in a development extractant DS5846, which is a formulation of a novel class of reagent currently under investigation by Zeneca Specialties. The properties of this reagent are described with respect to selectivity, isotherms and how they compare with phosphoric and thiophosphoric acids that have been previously proposed for the extraction of zinc. For the recovery of zinc from concentrated chloride solutions a solvating or neutral donor type of extractant is required which again should ideally have high selectivity for zinc over a number of other metals, particularly iron. The new reagent Acorga ZNX50 is based on a formulation of a novel class of compound developed specifically for this purpose. It gives efficient extraction of zinc from feed solutions that are four to six molar in chloride ion but is readily stripped at 60 - 65¦C using either water, dilute brine, or spent zinc chloride electrolyte. The reagent is highly selective for zinc over iron and a range of other metals including cadmium, arsenic, lead and magnesium.

Jan 1, 1993

The Electricity Commission of New South Wales is the largest single user of coal in the State, burning some 8.5 million tonnes during 1972-73. This coal is converted in power stations into electrical energy which is then distributed throughout the State and supplies the electricity requirements of 99% of users in New South Wales. Barring any major change in energy consumption growth, or change in the current mode of power generation, the consumption of coal by the Commission is estimated to have a doubling period of from 10 to 14 years. Coal is used to generate up to 87% of the electrical energy sold, the balance being obtained from hydro gen- eration, mainly in the Snowy Mountains, and, to a much smaller extent, from diesel generation at Broken Hill. A very large proportion of this coal is consumed in pulverised fuel boilers in the Commission's power stations, located mainly on the various coalfields of New South Wales. Coal supplies are drawn from a large number of mines, some owned or controlled by the Commission, others being owned by public companies. The coals used, originate from at least ten different seams located in the northern, western and southern coalfields of New South Wales, and therefore have a wide variety of properties and characteristics. To ensure an adequate fuel supply well into the future, the Commission is continuously engaged in coal prospecting in various areas. Bore data is used in assessing coal reserves and is also the basis of economic studies for their possible future exploitation. The numerous coal cores obtained during drilling operations are carefully tested and analysed in order to assess their suitability for power generation purposes, and to provide data for the design and operation of boiler plant using these coals. In existing power stations coal deliveries are systematically sampled and tested for quality control, for pay- ment of purchases based upon certain quality parameters and to check for any possible plant operating or main- tenance problems that might arise from the use of the particular coals. Almost all the properties of coal that may be necessary in the Commission's work are obtained using standard testing procedures carried out by the Commission's chemists, in laboratories located at the various power stations or at its Central Laboratories in Pyrmont. Where specialised analysis or testing is required, this is ordinarily undertaken on a commercial basis by outside bodies, expert in the particular field. The significance of the various properties of the coal for use in p.f. boilers is outlined and related to their proper application to boiler plant design and to plant operation.

Jan 1, 1975

Fundamentally, the development of a minerals processing project is a very risky enterprise. Most mining companies do not have sufficient financial resources to undertake significant projects and hence have to undertake external financing. Indeed, most resource companies understand the high risks associated in developing their projects and therefore, wish others to share in the risks. Most merchant banks, on the other hand, are searching to find suitable resource development projects in which they can share their risk and also share a large proportion of the ærewardÆ. To resolve this issue a Feasibility Study is completed by the project developer and assessed by the merchant bank often with the assistance of an æIndependent EngineerÆ (IE) who assesses the validity of the development proposal. Generally speaking, unless the Feasibility Study has been prepared improperly, the technical aspects of most mineral projects are reasonable. Some of the more recent so-called æhigh techÆ projects are exceptions but these still do not present major technical mistakes or misconceptions. p> In most cases, the major criticism the IE presents after discussing the validity of the ore reserves focuses on production plans, capital costs and operating costs. Of these three parameters the production plan is one in which owners tend to be far too optimistic and financiers, until recent years, have not taken a position. The most critical part of the development plan is the ability of the project, as designed, to attain the required design performance and, even more critically, the ability of the owner to reach and sustain this level of production. In short, how long should the æramp-upÆ be planned to take? Financiers have become more circumspect and are now taking a more conservative view of new project performance by factoring into their project financial analysis longer periods of ramp-up, higher capital costs and higher operating costs. However, these financiers can also be part of the æproblemÆ if they become overly conservative. This paper attempts, by way of case studies available from the public arena, to identify the pitfalls the project owners can meet and possible means of avoiding the disastrous financial conundrum whereby the second or third æownerÆ is the company that reaps the rewards initially extrapolated. The paper briefly discusses this problem by referring to projects known to the author and including previous studies that looked at this very problem. It is emphasised that many of the opinions stated within the paper are strictly those of the author and could be subject of much debate.

Jan 1, 2003