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Direct measurements of liquid contents in foams obtained from (x-terpineol a-T) n-pentanol (AmOH) and n-octanol (OctOH) solutions were carried out using a column of special construction. The column enables to determine the liquid contents by cutting-off foam layers of various thicknesses and measuring amounts of liquid contained in the cut-off foam layers. Liquid contents in the whole foam layers, in the top foam layers of thickness 14 ¦ 1 mm and distribution of water alongside foam height were determined. The average liquid contents were high and reached a level of 30, 35 and 40 per cent for foams formed from cc-T, AmOH and OctOH solutions, respectively. The liquid contents in the top foam layers were much smaller. They were of an order of five per cent for a-T and AmOH and 12 per cent for OctOH. On the basis of the liquid distribution alongside foam height some practical conclusions concerning possibilities of minimising amounts of liquid contained in the collected concentrate are presented.

Jan 1, 1993

Nippon Telegraph and Telephone Corporation (NTT) developed a new construction method (the M-2 method) that could drive a long, curved small-diameter tunnel (4>1.2 m) without human labor. This method employs the cast-in-place lining system, using quick-hardening resin mortar, so that it can greatly increase the driving distance and execute both long distance driving (longer than 500 m) and curved driving with a radius of curvature of 200 m. Furthermore, the central control system with a mini-computer in a vertical shaft enables all work in a small-diameter tunnel, i.e., excavation, lining, and transportation of excavated soil and materials, without human labor. NTT conducted several field tests and a Practical Application Test (P.A.T.) under a public road bearing great traffic. Good results were obtained and the practicability of this method was demonstrated. Therefore, from now on, NTT will apply this ii-2 method to drive telecommunication cable tunnels in urban areas. This method will also be applicable to constructing public tunnel facilities, such as for water, sewage, electricity and gas.

Jan 1, 1987

Studies of the minor element content of sphalerite and pyrite have generally been concerned with individual deposits or localised groups of deposits. However, a small number of more regional investigations have been reported (e.g. Davidson, 1962; Groves and Loftus-Hills, 1968; Loftus-Hills and Solomon, 1967; Price, 1972; Bralia et al., 1979) which have used the minor element abundances to evaluate the petrogenesis and geological affinities of diverse ore types and to fingerprint metallogenic provinces (e.g. Ivanov, 1964). While generally held to be valid, some doubts have nevertheless been cast on the reliability of minor element abundances in petrogenetic deliberations (e.g. Mercer, This paper reports on the distribution of he. Cpl and Mn in sphalerite, and Co, Ni and Mn in pyrite, from sulphide mineral occurrences within the "eugeosynclinal" prism comprising the New England Region of north-eastern New South Wales and south- eastern Queensland (Fig. 1). The occurrences examined are representative of two well-defined styles of mineralisation (Stanton et al., 1974) within which clear patterns of ore-rock distribution give rise to specific metal associations (Herbert. 1977).

Jan 1, 1987

By Kyle J. H, Misra V. N

The recovery of copper from a flotation concentrate using ferric chloride leaching has been investigated. The concentrate from the Horseshoe Lights Mine near Meekatharra, Western Australia, contained mainly digenite, pyrite and talc minerals. Over 95 per cent of the copper was dissolved in an acidified ferric chloride solution in one hour at 70¦C and ambient pressure. The pyrite and gangue minerals did not react. The leaching process was found to consist of four separate reactions and the shrinking core model was used to describe the process. The effects of particle size, temperature, concentration of FeCls, stirring speed and leaching time on metal dissolution were investigated. The apparent activation energies in the second, third and fourth reactions of leaching of copper concentrate were 28.8, 11.9 and 53.9 kJ/mol, respectively. The optimum conditions for the leaching of copper from the concentrate, within the range of conditions studied were: particle size 63 per cent minus 38 ¦m, a ferric chloride concentration of 1.48 M, a temperature of 70¦C, and a stirring speed of 800 rpm.

Jan 1, 1991

A continuing demand for exploration geologists, at a level well below that during the mineral "boom" of 1967-1971, will be highly selective. The exploration geologist must be more than the traditional geologist whose primary function has been detailed description of rocks and their recording on a geological map. He must be capable of 1. reconnaissance recognition of favour= able environments for ore mineral concentrations, 2. application and interpretation of techniques of search and evaluation, and 3. evaluation of economic, technological and political interactions that combine to make an ore deposit. Universities have an obligation to their students to provide them with a saleable skill, to educate them so that they can cope profess- ionally and socially, and to provide a mechan- ism by which education can be a continuing process. Humanistic and social studies, develop ment of communication skills, understanding of economic and management principles, ability to assess rock properties that influence extract- ion and beneficiation, and a knowledge of technological developments and limitations are necessary parts of the exploration geologist's

Jan 1, 1974

IA )W cli',hla cn1cnt fault 7011(-s 0 11Cn cOrnprise a nrc,II of inrcIIinkc(I faults, cvtcnqonal shear and In]rck cxtcnsir>nal vein- fractures. Hydrothermal mineralisation localised around such structures testifies to their importance as conduits for fluid flow. Mesh structures are most commonly developed in situations approximating bulk pure shear but predominantly simple shear meshes also occur. They appear to develop through the infiltration of pressurised fluids into a heterogeneous rock mass with varying material properties. In some circumstances, mesh formation appears to be a precursor to the development of major throughgoing faults. Mesh formation generally involves hydrofracturing (Pf> a3, at least locally) and the condition Pf - 63 to be maintained for the mesh to remain a high permeability conduit, requiring fluid overpressuring at other than shallow depths in extensional-transtensional regimes. The volumetric character of earthquake swarm activity appears to result from the passage of hydrothermal fluids through mesh structures, representing a form of distributed fault-valve action along suprahydrostatic hydraulic gradients arising from magmatic intrusion, compaction overpressuring, metamorphic dewatering, etc.

Jan 1, 1995

By W J. Bruckard, M J. Vaisey

Ashton Mining Ltd (Ashton) and Lynas Corporation Ltd (Lynas) have established a joint venture partnership to develop and process rare earth oxides (REOÆs) from the Mt Weld deposit in Western Australia. The deposit is located in a deeply weathered volcanic carbonatite structure that contains approximately two million tonnes of rare earth oxides at an average grade of 20 per cent REO. As part of the 1999/2000 feasibility study currently being undertaken by Ashton and Lynas, an extensive laboratory beneficiation test program was undertaken by CSIRO Minerals on diamond drill core samples from the Mt Weld orebody. The aim was to develop a flowsheet, based primarily on flotation that could be confirmed at pilot plant scale. The target concentrate grade was 52 per cent REO with a REO recovery of at least 35 per cent. Two main ore types, nominated as CZ (limonitic siltstone) and LI (nodular limonitic ironstone) were identified for separate laboratory evaluation. The lower grade LI-type ore overlies the higher grade and more extensive CZ-type ore. The majority of the work was conducted on the CZ ore; only limited testing of the LI ore was undertaken. The latter contains a clay component that will present difficulties in processing due to its poor filtering and settling characteristics. Two processing options for treatment of CZ ore were presented for pilot plant consideration. One (Option 1) includes gravity concentration prior to flotation while the other (Option 2) includes finer grinding and four stages of open circuit cleaning. The benefits of each scheme are discussed in detail. Neither has been fully optimised and the definitive flowsheet is most likely to be a hybrid of the two schemes, incorporating the best features of each. Typical metallurgical results from each scheme were final concentrate grades of 49 per cent REO, 10-16 per cent Fe2O3 and 3.2-4.8 per cent CaO. REO recoveries were above 35 per cent in both cases. Features of the processes developed by CSIRO were the inclusion of a gravity concentration stage prior to flotation, a modified flotation reagent scheme, a change in the method of feed preparation, and incorporation of magnetic separation to upgrade final flotation concentrates. Metallurgical results obtained from flotation tests on LI ore were much poorer than those obtained for CZ ore. The best rougher/scavenger result, with the limited amount of testing so far completed, was an REO grade of 32 per cent with a recovery of 28 per cent. Further work is needed to develop a viable processing route for this ore type.

Jan 1, 2000

A new gold cyanidation assay technique has been proven for a range of ores, where it has demonstrated benefits of being: ò Accurate (reduces the typical range of assay scatter almost three-fold, and avoids common fire assay interferences). ò Fast (under one hour, providing a quantum shift in space requirement and equipment utilisation each bottle roll can provide over 15 determinations per day). ò Simple (suitable for field application, no high skill or experience requirements). ò Profitable (reduced drilling, drilling time, ore misclassification and/or double handling after mining). ò Cheap (far lower cost than Fire or Aqua Regia Assays, ie LeachWELLTM costs around 10ó for a 20 g sample at 50 per cent solids, proportionately greater for larger samples eg $1 for 200 g samples which address most nugget effect problems). ò Relevant (distinguishes economic. ie cyanide-extractable gold, the only unequivocal value to use for true resource definition, mine planning and grade control). ò Robust (rough bucket chemistry of samplers does not seriously affect results). A new development is the provision of LeachWELLTM CAS, a solution, which also contains optimum cyanide and stabiliser concentrations; and offers greater: ò Safety (the dilute cyanide solutions are stabilised to avoid generating toxic HCN vapours; and present no cyanide dust or flying chips to deal with). ò Ease (a single squirt dispenses precise quantities of the entire reagent suite, avoiding multiple weighings). ò Reduction in total reagent costs. An inter-laboratory comparison on nine ore types, using two assay techniques (Fire Assay and Aqua Regia) involving 45 laboratories, has demonstrated that LeachWELLTM 60X assays for the samples tested, were equally or more reliable than typical Fire or Aqua Regia Assays. There is particularly great benefit in applying LeachWELLTM 60X assays to richer ores in comparison to Aqua Regia assays, which appear to be prone to very significant underassaying depending on individual analyst skills, and which often suffer greater nugget effects. The major impacts for heap leaching are that: ò Using the LeachWELLTM assay technique for grade control, ore of marginal grade is not accidentally discarded to waste or visa versa due to the scatter of assay data inherent in typical small charge (25 g) fire assays. ò LeachWELLTM assays very importantly report only that gold which will be recoverable in the heap, viz that which is liberated - an important factor where the coarse crush often leaves ten per cent of gold unliberated. These impacts are of particular importance for many ores (ie most of those recovered from open pit operations and at the margins of rich ore subject to slurry processing) subjected to heap leaching as the oxide zones often present major problems of nugget effects as well as mineralisation that cannot be accurately assigned through interpretation of clear structural controls.

Jan 1, 1994

By P Licence, M Stevens, P Atkinson, J L. Mauk

The Rahu and Ascot epithermal Au-Ag prospects are two km north of the Karangahake deposit in the Hauraki Goldfield, and all occur in an approximately 6 km2 north-south elongated magnetic æquietÆ zone. The prospects are hosted by andesitic flows, lithic-crystal tuffs and local sedimentary rocks; local silica sinter occurs at Ascot. These host rocks are intensely altered by quartz, adularia, albite, chlorite, pyrite, illite, interstratified illite-smectite and calcite. Several minerals show zoned distributions, such as widespread adularia that overlies more localised and deeper albite. Illite is the only dioctahedral phyllosilicate in several drill holes, but elsewhere it forms a shallow carapace underlain by interstratified illite-smectite with 50 to 90 percent illite. Veins are common with quartz dominating near the surface and calcite becoming more abundant at depth. Local laumontite-calcite veins also occur. Pseudosecondary and secondary fluid inclusions in quartz and calcite at Rahu and Ascot have homogenisation temperatures (Th) that range from 162¦ to 268¦C and 128¦ to 243¦C, respectively. Most inclusions have trapped dilute fluids with salinities of less than 1.9 weight percent NaCl equivalent, but 20 percent of the inclusions have salinities of 2.0 to 10.5 weight percent NaCl equivalent. Calculated paleowater table elevations suggest that the water table at Rahu occurred at 440 m above sea level (asl). At Ascot, silica sinter formed from discharging geothermal fluids (=100¦C) at the paleosurface at ~135 m asl. However the sinter also contains comb quartz with pseudosecondary inclusions that have a Th average of 229¦C, corresponding to a paleowater table at ~430 m asl during inclusion entrapment. Taken together, these results suggest that the Ascot sinter was buried by approximately 300 m of cover material during a major volcanic or tectonic event, and that subsequent hydrothermal fluids formed veins and alteration at Rahu and deposited comb quartz at Ascot.

Jan 1, 2007

By S Ellis, J McMullen, M Habner

A new leaching regime was implemented at Kalgoorlie Consolidated Gold MinesÆ (KCGM) Gidji plant on a cyanide leach of ultra-fine ground sulfide concentrate. High lime concentration and lead nitrate were used to overcome the refractoriness of the fine sulfides (high reactivity of fine sulfides with the presence of gold tellurides). The leach pretreatment stage was shown to be detrimental and was shut down. Lime titration was introduced as part of the process control. It was shown that the leach could be performed at low dissolved oxygen, (ie 3 - 5 ppm) to avoid sulfide oxidation and the associated high cyanide and lime consumption. Plant results showed reduced cyanide consumption (40 per cent) and an accelerated leaching rate, whereby most of the gold was dissolved in 24 hours compared to 72 hours used previously. Similar work was performed at KCGMÆs Fimiston plant, which leached a combined ultra-fine ground sulfide concentrate along with the slimes fraction from hydrocycloning the bulk flotation concentrate. The use of lead nitrate, high lime and carbon-in-leach generated a faster leaching rate and minimised adsorption of gold on fine pyrite particles. The grind, previously controlled with a P80, was replaced by a P98 to give a better operational indication of top size hydrocyclone performance. Before the implementation of the changes, average leach residues for slimes and UFG were respectively 4.0 g/t and 5.5 g/t. With the modifications to the leaching regime, average leach residues were lower by 1.5 - 2.0 g/t. Whilst significant improvements have been demonstrated, further work is progressing at KCGM to further reduce ultra-fine concentrate leach tail grades, and also to improve the understanding of concentrate refractoriness and the factors responsible for differences between laboratory and plant leach results. Estimated profits since the implementation of the new approach are in the range of $A3.5 million per year.

Jan 1, 2005

By V S. T Ciminelli

This work aimed at evaluating the acid rock drainage (ARD) potential of five samples of a Brazilian gold sulfide mine in a decommissioning process. Three of the most traditional ARD prediction techniques, Standard Acid Base Accounting (Std ABA), Modified Acid Base Accounting (Modified ABA), as static methods, and humidity cells as a kinetic method were tested. Both the Std ABA and the Modified ABA results indicated an acid generation potential for some of the studied samples. In the other cases the results were inconclusive. The five samples were then submitted to humidity cell tests during 27 to 30 weeks. At the end of the period, the leachate pH was stabilised in a range from 3.4 - 4.1 (four samples) to 5 (one sample). The presence of relatively high levels of sulfate and dissolved metals was detected throughout the operation, specifically, Ni and Zn, and less frequently, As. These findings are discussed with the help of the samples mineralogy and Eh-pH diagrams. The results obtained in this investigation indicated potential environmental risks of this mining area. In this particular case, the kinetic tests were shown to be essential to the identification of the samples ARD potential.

Jan 1, 2003

The T33-34 blast, the first major blast at Mount Isa Mines, broke 1,360,000 tonnes of copper ore using 349 tonnes of explosives. The original concept of firing T33, T34, T35 and U37 separately was changed during the life of the project due to adverse ground conditions. These conditions also caused the stoping method to be modified after the firing of the stopes had started. To stabilise the two stopes T33 and T34 where large fall off was occurring, rib pillars were left in the original slot. This caused the blast design to be more complex as these rib pillars had to "lead" the blast. The implication of such a large blast on the underground and surface environment was thoroughly investigated and precautions taken. The charging time of the blast was six weeks. Hook-up was started four days before the blast and continued until two hours prior to the blast. Detailed scheduling was carried out of other work that had to be done in the 24 hours before the blast to ensure that there were no delays in the firing.

Jan 1, 1977

By O Ostrovski

Direct iron smelting reduction has been examined using a modified 50 kg induction furnace. The furnace was equipped with continuous raw material feeding, a top oxygen lance, argon bottom bubbling, and sensors for measurement of gas and bath temperatures, gas composition and other parameters. In the base case, iron ore was charged at the rate of 1.5 kg/h, coke flowrate was adjusted to maintain a char fraction in the slag at 30 wt per cent, the oxygen flowrate was 100 Nl/min and the post-combustion ratio (PCR) was 60 per cent. The PCR was varied from 45 to 75 per cent by changing the lance height relative to the bath surface. Smelting reduction of Whaleback iron ore was investigated at different post-combustion ratios and iron ore feedrates using coke and coals with different volatile matter. The rate of iron oxide reduction increased with increasing iron oxide concentration in the slag, with a slight deviation from a linear function. The post-combustion ratio affected iron ore reduction and partitioning of elements between phases in the smelting reduction furnace. The high volatile coal investigated was found to be unsuitable for iron smelting reduction in the laboratory furnace.

Jan 1, 2000

By Wright S, Jorgensen F. R A

The reaction between lead sulphate and a number of base metal sulphides (galena, sphal- erite, chalcopyrite and synthetic pyrrhotite was studied as a means of desulphurizing lead sulphate containing fumes and residues to levels acceptable as ISF feed. The reactants were mixed with 50% zinc oxide to act as a "basket" to contain the molten products of the reaction and pressed into small briquettes 20 mm in diam. Factors such as temperature, time and the composition of the atmosphere were investigated. Using the PbS/PbS04 system as an example it was shown that the reaction became rapid above 800¦C and that both sulphur dioxide and oxygen in the atmosphere were detrimental to achieving low sulphur in the product. In an inert atmosphere all the sulphides when roasted for 1 h at 1000¦C gave results corresponding to the elimination of more than 95% of the sulphur in the reactants. Leach plant residues were considered for use as the "basket" material. It was shown that the residual sulphide and sulphate components of the residues underwent similar reactions to those experienced by the pure systems.

Jan 1, 1987

By M Zhu, J Safarian

Understanding the slag composition-structure-property relationship remains challenging due to the amorphous nature of slag melts and the difficulty of high-temperature experimental investigations. Since the CaO-Al2O3-SiO2 slag system is maybe the most important slag system for various hightemperature metallurgical process, the present research employed high-throughput molecular dynamics (MD) simulations on the entire composition region of the CaO-Al2O3-SiO2 slag system to determine physical properties like molar volume, density and diffusivity, structural properties such as distribution of different oxygen types for each simulated slag melt. The obtained slag properties result from the simulations displayed good consistency with experimental data, underscoring the reliability of the MD simulation as a novel tool to predict reasonable slag properties and to study slag behaviour from atomistic insight. The simulated results demonstrate a near-linear relationship between molar volume changes and composition, highlighting the predominant role of oxygen atoms, which is attributed to the significantly larger size of oxygen compared to other cations in the slag melts. This disparity in size leads to oxygen ions occupying most of the space in slags, thereby playing a major role in determining the overall molar volume. The simulation results of slag density revealed a decreasing trend with increased SiO2 content and identified a local minimum at constant SiO2 when Al2O3 content increases, also consistent with established slag density models. Furthermore, the simulated oxygen self-diffusion coefficients reveal a strong composition-structureproperty relationship, where an increase in CaO content enhances slag diffusivity, while a minimum in diffusivity is observed near a CaO/Al2O3 ratio of unity. The diffusivity minimum also corresponding to the local viscosity maxima, indicating effective simulation of charge compensation effects and dynamic characteristics of slag melts. The simulated distribution of oxygen types in slag — bridging oxygen (BOs), non-bridging oxygen (NBOs), free oxygen (FOs) and tricluster oxygen (TOs) — reveals a composition-dependent pattern: BOs peak near CaO/Al2O3 = 1, NBOs and FOs are abundant in CaO-rich areas with NBOs peaking near the Ca3SiO5 composition, FOs in the CaO corner, and TOs predominantly in the Al2O3 corner, highlighting the complex interplay of network formers and modifiers for the charge compensation effects on the slag’s structural properties. The successful application of MD simulations in this study paves the way for a powerful approach to explore the complex relationships in multicomponent slag systems and bridges the gap between mathematics modelling and experimental observations in high-temperature metallurgical processes.

Aug 21, 2024

By J Manuel, N Ware

Ideal iron ore sinter is produced by optimising the balance between coarse (+2 mm), stable ore nuclei and a reactive matrix consisting largely of fluxed -1 mm fines. The authors have previously published work evaluating the matrix melting behaviour of a number of fine ores, using techniques developed by CSIRO. The focus of that work was to demonstrate the contribution of the matrix to overall sinter quality, with the results showing a direct correlation between melting point, matrix porosity and strength. CSIRO has also developed a complementary test to study the behaviour of coarse ore nucleus particles in contact with the sinter matrix and provide an evaluation of their reaction and assimilation properties. This test provides valuable fundamental information that can be used to better understand the role of specific ore types and optimise blend composition.The CSIRO laboratory-scale assimilation test is carried out by embedding particles of ores of specific sizes into a fluxed matrix of known sinter blend composition, using real ores and fluxes, in a laboratory test that directly relates to pot-grate and plant sintering conditions. Samples are fired under controlled conditions in a tube furnace using a standard heating profile and a constant, low-oxygen potential atmosphere. These conditions simulate the actual sintering process while removing the effects of the many locally-variable parameters (eg temperature, oxidation potential and localised chemistry) that exist in the actual process, meaning that trends in behaviour can be more clearly seen. Designed as a compromise between the industry-standard pot-grate sintering test and a more fundamental test using chemical reagents, the aim is to isolate and evaluate the nucleus behaviour of an individual blend component to allow an improved understanding of the role that an individual ore plays in a specific blend. Measured cross-sections of the resulting fired samples provide a semi-quantitative measure of reactivity, while accompanying photomicrographs allow a qualitative assessment of assimilation properties (eg volume and distribution of secondary porosity generated during reaction).In this paper, to complement the previous work published on melting behaviour, nucleus particles were selected from hematite, hematite-goethite and goethite ores and embedded in a simulated plant sinter matrix. Drilled core particles of two sizes were tested at fixed temperatures simulating low- (<1300°C) and high-temperature (>1300°C) sintering conditions. The resulting assimilation properties are compared and discussed as a function of particle size, temperature and ore/sinter mineralogy. The implications for the contribution of each component to the prediction of blend sinter quality are also discussed.CITATION:Ware, N and Manuel, J, 2015. Fundamental nucleus assimilation behaviour of hematite- and goethite-containing ores in iron ore sintering, in Proceedings Iron Ore 2015, pp 77–82 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By M Skead, H Michael, R Chase, A Davidson

Ashanti Goldfields CompanyÆs (AshantiÆs) most recent development is the Geita Project in the Lake Victoria Goldfield of NW Tanzania. It comprises six distinct gold deposits within two mineralised trends: Nyankanga, Lone Cone, Geita Hill (Geita trend) and Matandani, Kukuluma and Area 3 West (Kukuluma trend). Between 1936 and 1966, the Geita area produced an estimated 5.5 M tonnes of ore from five underground deposits at an average recovered grade of 5.3 g/t. The discovery of Nyankanga was made by Ashanti in 1996 from RAB drilling of a +100 ppb Au anomaly in laterite soil above ferricrete cover. The Kukuluma trend was discovered by SAMAX Resources (taken over by Ashanti in 1998) between 1996 and 1997 from RC drilling of a +80 ppb Au laterite soil anomaly. Oxide facies banded iron formation (BIF) of the Upper Nyanzian System forms prominent topographic ridges and is the dominant host lithology for gold mineralisation. The ridges are flanked by felsic tuff units which lie stratigraphically above. Syn-genetic lamprophyre (vogesite) and post-genetic felsic dykes are also seen in the Geita trend. At Nyankanga a granodiorite-diorite suite has intruded as dykes and sills into BIF from a plug centred on a WNW fault and has produced wallrock alteration assemblages similar to those seen in Cu-porphyry systems (potassic, propylitic, etc) as well as sulphide and gold deposition. However, the main ore-grade mineralisation (as in the rest of the Geita trend) is controlled by low-to-medium angle structures dipping north-west and north with numerous splays in the hanging wall. High-grade shoots at Nyankanga plunge to the west and to the north-west, being lithologically and fault-controlled respectively. At Geita Hill the higher grade shoots are controlled by fold noses in BIF dipping to the north. Free gold occurs within fractures of sulphides. The deposits of Kukuluma and Matandani occur on an ancient (possibly Cretaceous) erosional plateau and are related to two WNW to NW trending shear zones, discordant to the local strike of the BIF sequence. Complete oxidation occurs up to 120 m depth and there is a strong supergene component to the gold distribution. Primary gold mineralisation is divisible into two distinct assemblages: arsenopyriteûpyrrhotiteûmagnetiteûchlorite associated with brecciated cherts and pyrite associated with shaley banded ironstones. Area 3W is most likely the fault-displaced extension of the same mineralised trend and comprises thin, intermittent zones of high-grade gold mineralisation occurring adjacent to a ridge in banded cherts and minor BIF. Current resources (measured, indicated and inferred) for the project are estimated at 89.24 Mt at an average grade of 4.17 g/t (11.97 Moz of contained gold) at a 1.0 g/t cut-off grade. Of this 9.83 Moz is within the Geita trend and 2.14 Moz is within the Kukuluma trend. A resource of 2.74 Moz is from an inferred resource at depth at Nyankanga where the viability of an underground operation is being studied. Proven and probable reserves are 49.77 Mt @ 3.46 g/t (5.54 Moz) based on a $325/oz gold price providing a mine life of ten years at an average production rate of 500 000 oz per year. Reserve estimation was carried out using mineable polygons to select resource blocks based on a series of mining and costs parameters. In addition key cost parameters were tested to provide sensitivities for pit optimisations and pathfinders for negotiations with suppliers and contractors. Construction began in March 1999 and first gold is expected to be poured in the second-quarter 2000.

Jan 1, 2000

Good morning ladies and gentlemen, and may I extend a particularly warm welcome to our shores to those of you who have travelled from overseas to attend this conference. I trust that you will find your visit to New Zealand both rewarding and enjoyable.   This is now my fourth appearance at a New Zealand minerals conference, the second of which has been jointly co-hosted by AusIMM and Crown Minerals. Since taking on the role of Associate Minister of Energy, I have had the opportunity to meet most of the industry players, and to visit a good number of the major mining operations. Through all of this I have become increasingly aware of not only how dynamic the minerals and coal industry is but also the array of factors that affect investment decisions in this business. Many of these factors, such as commodity prices and availability of investment funds, are determined primarily on the international stage. Others however, such as the allocation, regulatory and the fiscal regimes for minerals, are closer to home and are directly influenced by government policy decisions. I will talk more on what the government proposes in this area shortly.   But first, I would like to commend you all on your industryÆs continuing contribution to New Zealand Æs economy. 2004 has seen New Zealand Æs mineral and coal output top $1 billion for the second consecutive year. To put this achievement in perspective, it represents a staggering 47 per cent increase over the national output value recorded just five years ago. Notable contributions over the year came from gold which increased production to 10.2 tonnes (up nine per cent), ironsand production at 2.3 Mt (a 20 per cent increase), and industrial rocks and building stones (up six per cent) with rock for building, cement and roading showing significant rises on the back of strong regional growth and infrastructure development. Many regions, in particular Auckland, Bay of Plenty, Wellington and Southland, recorded significant increased regional activity with supply often struggling to meet demand. By contrast the coal sector had a relatively steady year with production remaining stable at just over five million tonnes for the second year running. Coal exports however are now consistently exceeding two million tonnes annually, generating export earnings of over $200 million, and could well be higher with rail capacity improvements. Many countries throughout the world are now importing our high quality bituminous coal for steel manufacturing and other industrial uses.

Jan 1, 2005

By G Solly, S Parsons, P Leevers, S Ni

Hematite ore has been mined in the Southern Middleback Ranges as blast furnace feed for the steelworks at Whyalla, by what was then BHP, since 1991. With the floating of BHP Integrated Steel as OneSteel in 2001 and the subsequent loss of key mining personnel, breakdowns in grade control methods and procedures occurred. These were compounded by changes in orebody characteristics and grade distribution leading to poor reserve recovery and poor grade control, both which resulted in loss of control of ore specification and severe deviation from the mining schedule. As a result, a total review of grade control, mining method and scheduling procedures was undertaken as well as a review of contractor quality assurance. It was found that sampling methods were poor and unrepresentative, grade control blocking processes were outdated and did not reflect the geology of the deposit, the ore mining method was outdated and did not respect the geometry of the orebody, reconciliation methods were being completed globally and were not reflective of individual areas of the deposit, and that deviations from the mine plan were occurring due to access to high-grade areas outside the schedule in order to bring stockpiles back on grade. Based on these findings, updated sample splitter systems were fitted to drill hole rigs, double bench sampling was implemented (2 + 4 m samples as opposed to 1 + 8 m sample) and a three-dimensional cellular grade control system introduced to better reflect the geology and orientation of the deposit. At the same time ore mining was changed from front end loaders mining on 8 m benches to excavator mining on three lifts on 8 m benches; this aided selectivity and respected the geometry of the orebodies aiding cleaner recovery. Reconciliation is now completed on a blast-by-blast basis, which allows for quick and easy identification of poorly performing areas of the resource model that can then be incorporated back into resource model updates. Regular daily and weekly meetings are also held with emphasis on the mining schedule to ensure no deviation from the schedule in the long term. These improvements have resulted in a lift in reserve recovery both globally and locally associated at the same time with improvement in grade.

Jan 1, 2005

By M Noy, P Hawkes

Improvements in the drill and blast process can impact on loader performance in a number of ways. The main effects being improved loading cycle times, less non-productive cycles such as oversize removal, clean-up and face raking and better bucket fill factors. The impact of improved loading cycle times on the overall cost and productivity of the load and haul cycle is very much dependent on the loader truck match and the management of the haul fleet. Previous work has shown that a 20 per cent improvement in dig cycle times may only lead to three per cent or less improvement in the overall productivity of the load and haul process. However, improving the mass loaded per pass by five per cent will lead to a five per cent improvement in productivity for any given haul fleet situation, assuming sufficient truck capacity. Work by Orica Explosives at Lab ChrysotileÆs Thetford Mine, Quebec, Canada showed a relationship between muckpile fragmentation and truck payload. With this in mind, work was carried out to determine the effect differing muckpile fragmentation distributions can have on excavator bucket fill factors. This work was performed in the laboratory to a 1:100 scale. Nine different distributions were used. These were characterised by the Rosin Rammler parameters, xc the characteristic size, and n, the uniformity index. The values of xc and n used were modelled on those previously measured at a Sydney quarry. For the range of distributions measured, it was found that the fill factors were lowest at a low characteristic size and a high uniformity index and highest at a high characteristic size and a low uniformity index. In other words the best fill factors were achieved with a more dispersed muckpile distribution with a larger characteristic size. The difference in fill factors was 16 per cent indicating that controlling the drill and blast process to produce a specific muckpile distribution can have a significant impact on the overall cost-efficiency and productivity of the load and haul process. Field measurement by Orica Explosives at a quarry in Canada supports this conclusion showing improvements of up to seven per cent in average bucket load by either increasing the characteristic size or decreasing the uniformity index of the muckpile.

Jan 1, 1999