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By A Annor, L J. Chen

The production of paste backfill has typically been accomplished through the adaptation of existing technology. Thickeners and disc filters have several drawbacks when applied to production of tailings paste backfill. Personnel at CANMETÆs backfill research laboratory in Sudbury, Ontario, Canada have been carrying out research into and development of methods to produce paste backfill using total mill tailings since 1991. Using a silo for production of paste is not a new concept but has been refined to a marketable level. This technology is based on a fluidization principle that assists in settling the tailings to an homogenous state in a silo. Other developments include patented de-watering technology for use in the silo to speed the settling of the solids to a suitable solids content, optimisation of silo geometry, and instrumentation related to paste backfill production and transportation. The advantages to fluidisation technology are in reduced capital cost, reduced operating costs and in the environmental benefits of using full plant tailings over classified tailings or alluvial aggregates. Another major advantage is in the ability of this technology to be retrofitted to existing hydraulic backfill plants to further reduce capital cost. This paper details the development of the new technology as well as the associated cost savings over other technologies. It will also look at the research and development of unique components that make this technology work.

Jan 1, 1998

By A Sadri, S Smith

In the metal and mineral industry, different empirical techniques are used to estimate the remaining thickness of the lining within an operating furnace. The disadvantage and costly factor when using these methods is that production needs to be halted. Hatch has developed an accurate and reliable technique based on stress wave propagation principles to determine the condition of the remaining lining whilst the furnace remains online and operational.The technique involves test data collected and compiled to generate 3D imagery used to assist operators and maintenance planning personnel with valuable information showing areas of metal penetration in the lining, remaining lining thickness, exact location of voids, cracks, poor contacts and deterioration due to chemical attack. In this paper, the principals of the AU-E technique will be explained both theoretically and practically. In addition, the use of ultrasound technology for determining the condition of copper cooling elements will be briefly addressed.

Jan 1, 2006

Sample preparation for gold samples, where the distribution and size of gold particles is often irregular, usually demands that a relatively large sample weight is taken to produce a representative sample for analysis. Traditional sample preparation of gold samples incorporated multiple stages of coarse particle size reduction followed by subdivision of the sample prior to final fine pulverising. This technique developed because of the limited equipment options available to laboratories. As a result significant errors are introduced and the final sample size being pulverised is a small fraction of the original sample weight. Three significant developments have occurred in the last 25 years to significantly improve sample preparation of gold samples: 1.  large capacity fine pulverising,2.  fine jaw crushing, and3.  whole sample coarse pulverising. In the mid-1980s an Australian manufacturer developed large capacity pulverising bowls which had the ability to reduce samples with a mass of 1.5 - 3.5 kg to a nominal 0.075 mm product in approximately three to five minutes. Laboratories were able to produce a large homogeneous pulverised sample which significantly reduced the error in sampling. However, exploration and mine samples are usually 7 to 10 kg and the sample requires a subdivision stage following crushing to produce the 3.5 kg for the pulverising step, thus introducing potential for significant error. The development of fine jaw crushing allowed particle size reduction to -2 mm which reduced the subdivision error of the bulk sample prior to fine pulverising. Subdivision of the crushed sample at -2 mm still results in potential sampling errors, where the gold particles are not sufficiently liberated to be representatively sampled. Clearly the best option would be to reduce the particle size small enough to ensure a representative sample prior to final pulverising. Continuous mills offer a partial solution but are relatively slow and subject the sample to differential pulverising prior to subdivision. Clearly this does not provide the best solution for routine laboratory applications where high volumes of samples are to be processed. The recent development of a large capacity coarse pulverising system allows samples of up to 10 kg to be reduced to -0.25 mm prior to subdivision, and fine pulverised to -0.075 mm. Laboratories now have the necessary equipment to perform reliable large capacity sample preparation of gold samples. The exact choice of a sample preparation scheme may still depend on particular ore types but the need for multiple crushing and subdivision stages is redundant.

Jan 1, 2008

By A D. Pradana, D R. Pradana, Y H. Putra, Z Zulhan, T Hidayat

The high temperature processing of nickel laterite ore is currently dominated by the Rotary Kiln- Electric Furnace (RKEF) technology. The ratio of SiO2/MgO (S/M) in the ore is one of the critical parameters that determines the success of the RKEF process. Incompatible S/M ratios in the ore can cause aggressive chemical interaction between the slag and MgO refractory. The slag-refractory interaction at conditions relevant to nickel laterite ore smelting was investigated in the present study. Synthetic mixtures representing the slag compositions of nickel laterite ore smelting were prepared by mixing MgO, SiO2, Al2O3, CaO, Fe2O3, and Fe and heating the mixtures in steel containers at 1300°C under an argon atmosphere for 3 hrs. Pure MgO crucible were used as containment material in the melting process of the mixtures to represent the MgO refractory. The kinetics of the slagrefractory reaction were investigated using samples with selected initial S/M ratios of 2.0 and 3.0 (wt/wt) (with constant CaO = 3 wt per cent, Al2O3 = 5 wt per cent, FeO = 10 wt per cent) by performing melting at 1500°C in MgO crucibles for 5, 30, and 120 mins under an argon atmosphere. The effect of the S/M ratio in slag on the slag-refractory interaction was also investigated by performing melting of mixtures with different initial S/M ratios of 1.75, 2.0, 2.5, and 3.0 (wt/wt) in MgO crucibles at 1500°C for 120 mins under an argon atmosphere. After the melting process, each sample was quenched in water, mounted in resin, ground, polished, and coated. The interaction between the slags and MgO refractory was then evaluated by analysing the polished samples using a scanning electron microscope equipped with an energy dispersive spectroscopy (SEM-EDS). Thermodynamic simulations using FactSage™ 8.0 thermochemical software were performed based on the experimental slag compositions, refractory, and conditions to support the evaluation process. The results show that a higher S/M ratio in slag leads to a higher tendency for the refractory component to dissolve into the slag. A higher S/M ratio in slag also leads to an increasing extent of solid-state diffusion of iron oxide into the remaining refractory grains. The dissolution of MgO into the slag is rapid; the MgO concentration in the final slag increased by 2.4 wt per cent and 5.7 wt per cent at initial S/M ratios of 2.0 and 3.0 wt/wt, respectively, within 5 and 30 mins of the melting process.

Aug 21, 2024

HISTORY OF PROCESS DEVELOPMENT The manufacture of coke from coal commenced at the beginning of the 17th century by heaping coal into circular mounds in the open, covering with earth, clay or ash and slowly burning the coal in the mounds to produce a heap of coke. Naturally the process was slow in production rate with low coke recovery due to the loss of that part of the coal which was completely burnt to ash. As a natural consequence of this initial carbonisation method, a coke oven design was developed so that the coking chamber, which was constructed of brick, was circular in plan, cylindrical in section with a dome shaped roof - similar in shape to the original coal mounds. As the shape of the chamber was similar to the old dome shaped beehive, the ovens were called beehive ovens. The ovens were built in long single or double rows with retaining walls between each oven and as such were known as a battery of beehive coke ovens. In these early beehive ovens, coal was charged from hand wheeled tubs into a hole into the top of each oven, hand- levelled with a rake and the oven side opening sealed with bricks and clay except for some small holes in the opening to admit limited quantity of air for combustion. Coking commenced by the heat retained in the brickwork from the previous charge. The gases evolved from the coal, together with the admitted air, were partially burnt inside the oven to supply the heat for carbonisation and the surplus gases were burnt to waste at the top charging hole. The heat from this gaseous combustion inside the oven together with the reflected heat from the brickwork of the arched roof of the oven allowed carbonisation of the coal charge to proceed from its surface slowly downwards to the floor of the oven. At the end of the coking cycle (usually 4 or 5 days) when there was no more evolution of gases, as indicated by diminishing gas flame and a drop in oven temperature, the brickwork of the oven opening was removed. The coke was then quenched by inserting a pipe with spray holes to spray water over the coke charge in the oven after which the coke was raked from the oven. The work of charging and levelling the coal, bricking and removal of oven opening, quenching and raking out the coke all required manual labour. As these ovens only carbonised a 5 to 7 ton charge of coal per cycle of 4 to 5 days, to produce 3 to 41/2 ton of marketable coke, the labour required per ton of coke was high.

Jan 1, 1975

In the course of an investigation of the influence of bismuth on copper carried out some years ago, (1) the author found that phosphorus exerted a beneficial effect on bismuth-copper alloys of low bismuth and oxygen content by neutralizing the embrittling effect of the bismuth. In the intervening years, a further examination of bismuth-phosphorus-copper alloys has been canied out and the influence of other elements and combinations of elements on copper containing bismuth has been studied.The results recorded in the present paper are part of this investigation.MATERIALSAll the alloys were made from high-grade copper (KS.A. wire-bar) of the following composition: copper 99·95%, oxygen 0,040'%, iron 0·0015%, sulphur less than 0-001%, bismuth less than (j>'OOl%, other elements each less than O· 0010/<-. The bar was hot rolled and cold drawn to strip which was cut into short lengths and cleaned in nitric acid ready for melting.A piece of this 'strip was drawn to wire 0·08 in. diam., annealed at 5000 C, and measured for electrical conductivity as a check on purity. The result was 101·5% of the I.E'.C. standard.In making the alloys, phosphorus was added as 14% phosphor-copper, bismuth as 17% bismuth-copper, and the other constituents either as copper alloys or as elements of a high degree of purity.PREPARATIQN OF STRIPSThe charges were melted in plumbago crucibles (Morgan salamander brand) in an oil-fired furnace, Le;, in the manner commonly adopted in industrial practice. The copper was first melted under charcoal, well stirred with a chaned stick, and the necessary alloying elements added and stirred in with the stick. The weight of a charge varied according to requirements but generally was about 4 lb. The charges were poured into slightly tapered brass moulds heated to about 1000 C and dressed with a mixture of mineral oil and tallow. They were not only poured under charcoal through the burning mould dressing but also- through a ring of flame produced by burning kerosene on asbestos cord wrapped round the top of the mould. Experiments carried out with plain copper proved that by this means oxygen could be excluded from the stream of metal. The billets were 7·5 in. long x O·83 in. diam. x 1·3 lb. weight.The billets were cropped about 1 in. from the top, cleaned in nitric acid, and measured for density by the usual method of weighing in air and -in water. As these measurements were required only as a check on casting, technique a high degree of accuracy was not essential (accuracy = -+ 0·2%).

Jan 1, 1942

Limnological processes are often overlooked when trying to assess the long-term water quality of pit lakes. It is essential to deal with these aspects, since they determine the geochemical and redox conditions around the submerged reactive rocks. Underground mining at East Sullivan (northwestern Quebec, Canada) stopped in 1966. Two adjacent stopes, 91 and 106 m deep, were left open to the surface, with roughly circular apertures of about 7700 and 18 000 m2. Water filling of the two glory holes went on over three decades and overflow began in 1999. Surface water pHs values were near neutrality as late as 1994 (6.3 - 7.0) but had become acidic by 1999 (2.6 - 3.5). Limnological processes were investigated in late-1999 and early-2000, with temperature measurements at 1 m intervals and sampling of the water columns before and after the late fall density inversions. A strong chemocline-thermocline was observed at 15 m. Above the chemocline, the water had low pH (3.2), no alkalinity, and low Na, Ca and sulfate. Below the chemocline, the concentrations of these parameters and the pH increased, with pH at 5.5 - 6. Consistent with the low pH values present in the upper 15 m, the concentrations of Al, Cd, Cu, and Zn were high near surface, and decreased below the chemocline. Dissolved oxygen concentrations as well as Eh decreased with increasing depth. Alkalinity integrated over the entire water column below the chemocline was at least six times larger than the acidity inferred from Fe2+ and Fe3+ concentrations. The temperature survey showed that seasonal density inversions do occur in the water column. The inversions do not involve the entire water body, since the surface to depth ratio is too small for a complete overturn. Oxygen from surface waters is carried to the bottom where it oxidises sulfides, until its depletion. Residual Fe2+, stable under anoxic conditions, is displaced upward by the next dense water incursion. It eventually reaches the thermocline, but only after many successive incursions. There, its oxidation to Fe3+ allows hydrolysis, which consumes dissolved and wall-rock alkalinity. It is only in 1999 that alkalinity was finally exhausted and that the pH shifts to acid values. The low pHs in the surface waters allow a build up of Fe3+ and Al3+, and desorption of metals from colloids. Buffering is assumed to be controlled by wall-rock oxyhydroxides in surface waters, and by carbonate minerals in the bottom waters. Because the seasonal overturns are far from being complete, the large amounts of alkalinity acquired from the wall-rock in the bottom waters cannot balance the acidity of near surface waters. Hence the surface water is acidic.

Jan 1, 2003

By A Fareth

Constructed between 1885 and 1891, the New Croton Aqueduct (NCA) in New York State is a 50 km long (31 mi) brick-lined tunnel conveying water from the Croton watershed north of New York City to distribution systems in the Bronx and Manhattan. Except at two siphons, the aqueduct operates as an open channel conduit from the New Croton Reservoir downstream 39 km (25 mi) to the Jerome Park Reservoir in the Bronx. The remaining 11 km (6 mi) is operated as a pressurised conduit and includes a major siphon, of circular cross-section, under the Harlem River. The aqueduct has a flow capacity of approximately 1100 million litres of water per day (13 m3/s or 290 million gallons per day) and supplies on average ten per cent of New York CityÆs drinking water. The CityÆs water supply system is operated and maintained by the New York City Department of Environmental Protection (NYCDEP). This paper covers the work done during the three main stages of the program to rehabilitate the entire NCA: inspections, design and rehabilitation. It also includes some of the more interesting structural findings of the original construction as well as documents and drawings from the NYCDEP archives. The rehabilitation program has been underway in phases since 1993, with inspections scheduled to minimise outages, allowing this strategic aqueduct to remain in service during critical seasonal periods. The open channel flow section of the aqueduct in the north is horseshoe shaped, with inside dimensions of 4.1 m (13.6 ft) in width and 4 m (13.5 ft) in height. The pressurised section is 3.7 m (12.3 ft) in diameter, and includes a 90 m (330 ft) deep siphon under the Harlem River. A total of 163 million bricks were used in the construction, enough to build a 48 km (30 mi) wall around Manhattan Island, 3 m (10 ft) thick and 15 m (50 ft) high.Between 1993 and 1997 a series of in-tunnel investigations were performed in the open channel portion of the NCA, consisting of field inspection, non-destructive geophysical testing, and coring of the brick liner. Between November 2004 and September 2005, a major inspection program was conducted to assess the condition of the 11 km (6 mi) long pressurised section and all shafts, headhouses and blow-off structures along the entire 50 km (31 mi) alignment. Inspection methods included using an underwater remote operated vehicle (ROV) equipped with sonar and cameras to inspect the deep siphon, fibre optics examinations of probe holes drilled through and beyond the brick lining, and core holes and geophysical inspections to assess properties of the liner and behind-the-liner materials. A water pressure and test grouting program was conducted to assess methods and expected grout takes for the rehabilitation work in the pressurised sections. The rehabilitation program is anticipated to be completed in 2010.

Jan 1, 2008

By D Judkins N OÆLoughlin

"Basin Minerals Limited commenced exploring for conventional strandline hosted mineral sands (rutile, ilmenite and zircon) deposits in the Murray Basin in April 1997. This has resulted in the discovery of a major mineral sands field in the Douglas Project area, located southwest of Horsham in the Wimmera Region of Western Victoria. After the successful completion of a pre-feasibility study on these deposits in September 2000, the Douglas Project has advanced to bankable final feasibility study. This study should be completed by mid-2002 to allow a decision to commence production in 2003 at an expected rate of in excess of 500 000 tonnes of combined mineral sands products per annum. The field comprises several mineralised strandlines. The main ones of economic interest are from west to east, the four strandlines, Acapulco, Bondi, Bondi East and Echo. These strandlines at August, 2000 contain total Inferred Resources of 443 million tonnes at 4.7 per cent HM at a one per cent HM cut-off; ie 20.8 million tonnes of heavy mineral concentrate which in turn contains 11.3 million tonnes of ilmenite and leucoxene, 1.3 million tonnes of rutile and 1.6 million tonnes of zircon. The heavy mineral concentrate is coarse-grained with mean diameter grain size of the strandline deposits ranging from 130 to 160 microns in size. Major controls on the formation of mineral sands deposits in the Murray Basin are discussed with reference to the main mineral sand bearing, strand plain domains; the Central Strand Plain, the New South Wales Strand Plain and the West Wimmera Strand Plain. The Douglas Project deposits on the West Wimmera Strand Plain domain are characterised by wide zones of mineralisation with high-grade cores (>10 per cent HM), low stripping ratios, and a valuable economic heavy mineral suite comprising 60 to 80 per cent of heavy mineral concentrate. Constituent mineralogy of the heavy mineral concentrate of the overall deposits is; ilmenite (40 to 58 per cent), leucoxene (five to seven per cent), rutile (four to ten per cent) and zircon (four to 14 per cent). The Douglas Project Area appears to be endowed with significantly enhanced levels of heavy minerals compared to other areas of the Murray Basin due to its unique palaeo-geographical position. Long shore drift to the south of the basin has caused large quantities of heavy mineral to be trapped against the palaeo-cliffs and headlands of a rugged palaeo-coastline which includes in a major embayment, the Grampian Mountains, the Black Range and the Glenelg Complex. To the east, major faulting and uplift has brought the fine-grained WIM offshore deposits close to the surface and has resulted in erosional stripping of mineralised strandline deposits. This mineralisation has been recycled and reconcentrated within the younger strandlines to the west within the adjacent Douglas Project Area. These strandlines are thought to be younger than other mineralised strandlines discovered to-date in the Murray Basin. The Project Area strandlines are described; ie their morphology, geology and patterns of mineral distribution. The mineralogy of each of the strandlines is distinctive and variations are noted along strike and section."

Jan 1, 2001

By A Coulson, J Hussey

Noranda Inc is a Canadian mining and metallurgical company with its activities managed through businesses in copper and recycling, zinc, aluminum, magnesium and various metals through its interest in Falconbridge Ltd. Noranda operates three base metal mines, three copper smelters, one copper refinery, one zinc refinery, one lead smelter and refinery, three copper recycling and sampling facilities seven aluminum producing facilities and one magnesium production facility in the construction stage. The mining side of Noranda has a strong history of open stope, cut and fill and shrinkage mining. The 1999 Noranda Annual Report lists a 206 million tonne resource located at Porphyry Mountain in GaspT, QuTbec. Mont Porphyre is a large tonnage, low-grade porphyry style Cu-Mo deposit that was discovered in 1994. It is located at a depth between 1.0 and 1.7 km below surface and characterised by a very competent rock mass. The Mont Porphyre resource is located adjacent to the town of Murdochville, close to NorandaÆs Mines GaspT operations. Mines GaspT closed its open pit and underground operations in 1999 after approximately 45 years of mining, but still operates a copper smelter near Murdochville. The low-grade and great depth of the Mont Porphyre resource suggested that mine design could not be based on the conventional Noranda open stoping mining practice. This paper will review NorandaÆs approach to evaluating the Mont Porphyre deposit. The approach included two definition diamond drilling campaigns, a review of current caving practice, an in-house scope study, a geotechnical logging program for diamond drill core and participation in an international study on cave mining. One interesting component of NorandaÆs approach was the scope study that involved a team of multi-disciplinary personnel drawn together for a short period of time. The objective of the scope study was to determine what conditions would be necessary to mine the deposit and to assess the merits of continuing with further exploration work. A review of current worldwide underground bulk mining practices, and in particular an evaluation of the applicability of caving mining methods, were key components of the scope study. The scope study team concluded that derivatives of caving could be applied to the mining of Mont Porphyre, but additional diamond drilling was recommended to further define the deposit and improve the level of geotechnical information. Following the recommendations of the scope study team, a diamond drilling campaign was completed during the summer and fall of 1997. The 1997 drilling campaign had two secondary objectives, in addition to the primary objective of defining the limits of the orebody. The first objective was to improve the level of geotechnical information available for the characterisation of primary block cave fragmentation. The second objective was to improve the delineation of a lower grade intrusive core. A geomechanical investigation was conducted and included particular emphasis on geotechnical data collection and analysis. A review of the actual data collection process, including several trials of different core orientation techniques, will be included in the discussion.

Jan 1, 2000

By J M. F Clout

The low-grade clay-rich -1.0 mm fraction of fine iron ores have often proved difficult to upgrade and capital intensive to process when using traditional wet beneficiation technology including single or two stage hydrocyclone desliming, wet high intensity magnetic separation (WHIMS), spirals and final dewatering using fine screens. Very fine (<0.010 mm) clay-rich slimes simply follow the water split to the hydrocyclone underflow where it often impedes the subsequent processing performance of any downstream spirals and can result in troublesome direct dewatering of hydrocyclone underflow. The -1.0 mm fraction in goethitic iron ores is typically considered as waste. A new two-stage wet gravity separation flow sheet, referred to as the ædesandÆ process, has been designed to overcome these issues for the upgrading of low-grade clay-rich goethitic or hematite-quartz fine iron ores. The desand process is designed to treat the -1.0 mm range from -8.0 mm natural fines, to recover a -1.0 +0.038 mm higher-grade fraction by rejecting to tailings any low-grade clay-rich ultrafines and low density sand-size fine chert/quartz gangue and shale.The desand process consists of wet screens with a nominal 0.7-1.0 mm cut point which direct the undersize to hydrocyclones. The hydrocyclone underflow is then fed to upcurrent classifiers where the classifier middlings is sent to spirals and classifier and spirals concentrates are then dewatered using vacuum belt filters or fine screens. The three product upcurrent classifier is employed to further deslime the hydrocyclone underflow, producing firstly a coarse final concentrate, and then a deslimed middlings for feeding to spirals as well as a 0.100 mm clay and fine quartz or banded iron formation (BIF) sand in final overflow for feeding to the residue thickener. The upcurrent classifier middlings provides the spirals with a significantly cleaner feed material than traditional methods, with the end result that the spirals do not silt up with slimes and therefore can continue to run at their optimum. The upcurrent classifier and spirals form a complimentary pair, with the upcurrent classifier better able to send coarser, acceptable grade particles to concentrate and fine slimes to tailings whilst allowing the spiral to better reject coarser, subgrade low specific gravity particles. Test results for the Mayoko-Moussondji iron project, Republic of Congo, using upcurrent classifier-Wilfley Table (simulating spirals) gave acceptable mass recovery from run-of-mine (ROM) and Fe recovery from hydrocyclone underflow for HEMCOL (6.2 per cent mass recovery from ROM/43.1 per cent Fe recovery) and HEMBIF2 (5.1 per cent mass recovery from ROM/22.3 per cent Fe recovery) mineralisation types. The desand process also enables improved dewatering of -1.0 +0.038 mm product from a clay-rich plant feed.A number of new wet plants in the Pilbara use the desand flow sheet including FMGÆs Cloudbreak and their two Christmas Creek plants. Many potential new applications are being explored, with new iron ore plants in India as well as those being proposed for West Africa. Operational experience has highlighted the importance of providing clean process water to the upcurrent classifiers as well as ongoing design improvements including increased wear life, reduced maintenance and enabling a more steady feed rate to spirals.CITATION:Clout, J M F, 2013. Beneficiation of fine iron ores using the desand process, in Proceedings Iron Ore 2013 , pp 337-348 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

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 R C. Schodde

The future of any mining company depends on maintaining and growing its access to high-quality mineral resources. Often the true size of a given resource is not fully recognised until after mining has commenced. Also, the strong desire to increase production puts pressure to expand the resource base of the mine to get maximum value out of the capital employed. This, plus diminishing M&A opportunities and declining world-class greenfield discovery rates, makes brownfields exploration a high priority activity in any operating company. For the success and long-term survival of the company it is critical that this is done well. The primary aim of brownfields exploration is to find or acquire new deposits within economic transport distance of an existing mine that will add materially to shareholder value through: mine or capacity expansion, displacement of lower-margin ore from the mine plan, preserving and extending the mine life, and ensuring potential value is understood in order to maximise potential future mine development or exit. A number of strategic principles are important in any program designed to achieve these aims: understanding the life of mine (LOM) production profile and the related key drivers for exploration, understanding the dynamic nature of the brownfields search space, understanding and measuring depletion of the brownfields search space, understanding the potential of the near-mine region, investing in exploration technology to expand the search space, and investing in a basic geological understanding of the ore-environment. Examination of the LOM profile will highlight critical issues facing development plans for the mine including future grade and throughput issues, key technology requirements or innovations and scenarios for increased production. Exploration targets in the brownfields program should address these key issues. Often it is as æsimpleÆ as keeping mine development sufficiently far ahead of extraction to allow time for the various phases of exploration and resource assessment drilling. Because of the low additional capital costs required, expansion of the brownfields search space is strongly leveraged to: metal prices, development of new markets, incremental operating cost reductions, strategic changes in downstream requirements in integrated businesses, depletion of higher unit-value ores, and new technologies (both extraction and exploration). Therefore the brownfields search space is likely to incrementally expand through time and must be regularly reviewed. This is also important as any given search space will be progressively exhausted, resulting in smaller and higher cost discoveries. It is therefore important to monitor and critically review sustained trends of increasing brownfields discovery cost. The first wave of discoveries in any camp are usually empirically driven and do not need a strong geological understanding. However, an improved geological understanding is critical to making the next generation of hidden discoveries. To illustrate these points, a case study was carried out on the exploration and development history of the Kambalda nickel camp. Over a period of 40 years, more than 40 deposits were found containing over 1.7 million tonnes of nickel metal. Efficient mines need to maintain reserves for operating flexibility. Consequently, the company needs to clearly understand the global resource in order to maximise its full potential. In addition, understanding the full brownfields potential earlier in the development of a deposit rather than later, creates optionality through the flexibility that this understanding brings to future development choices. To ensure best practice in attaining this aim, the roles and responsibilities of the exploration team and the mine operation should be clearly understood. It is important for the explorers and the operators to have shared views about what needs to be done. Typically, mine operators would be accountable for identifying future resource requirements to support the life-of-mine production schedule whilst explorers would be accountable to ensure best-practice assessment of brownfields exploration opportunities. This would include an assessment of the global undiscovered endowment in the regional near-mine area. The balance between brownfield exploration drilling and resources to reserve conversion will need to be decided to ensure efficient operations are maintained. Both need to agree the optimum level of brownfields funding taking into account: urgency to replace or grow reserves, and quality of opportunities (relative to those elsewhere). The ultimate cost of ignoring the full brownfields potential of an operating asset can be significant, ranging from inefficient and poorly planned operations, lost opportunity cost and resultant value through under-utilised resources and, in the worst case, early exit from the asset at the bottom of the commodity cycle û leaving behind huge amounts of shareholder value.

Jan 1, 2006

By D Corley, J Moore, M Grant, A, W R

Resource estimates are the corner stone of technical and investment decision-making. Prior to mining, resource estimation uncertainty has the greatest potential to lead to poor investment decisions, despite a significant component of resource estimation uncertainty being unknowable at this critical stage in the mining cycle. This presents a conundrum to the resource geologist in terms of risk evaluation and resource classification. Ahead of mining, there is normally considerable focus on drill hole spacing analysis to determine the ‘optimal’ drill hole spacing, taking grade and geological continuity into account; cost-benefit analyses balance improved resource definition against cost; the law of diminishing returns and the exponential cost increases acting together as spacing is reduced. Typically, a lower limit to the drill hole spacing is identified. After completing the resource drilling to an agreed spacing, resource estimation is then undertaken. At this point it is common to attempt to bracket the resource estimation uncertainty. In many cases conditional simulation is used, based on a modelled, but largely assumed, variogram model. It is also assumed that the histogram of the available drill hole sample data is representative of the in-ground mineralisation. The purpose of this paper is not to diminish the importance of drill hole spacing and simulation studies, but rather to illuminate an important aspect of estimation uncertainty that, although previously recognised, is typically overlooked – the question of how representative the available data set is in characterising the true (but unknown) distribution of mineralisation. To do this, we lift the lid on one of OceanaGold’s former operations, the mined-out Globe Progress Mine, by resurrecting a high quality, close-spaced, reverse circulation (RC) grade control data set. The Globe Progress Mine is within the West Coast Region of New Zealand, it was closed and transitioned to rehabilitation in 2016 and is now known as the Reefton Restoration Project. The exhaustive Globe Progress grade control data was used to repeatedly ‘redrill’ the deposit by extracting 35 m × 35 m spaced subsets from the original 5 m × 5 m spaced grade control data. Utilising closely spaced grade control data removed the need for assumptions regarding short range continuity, which are necessary with forward-looking analyses that are based upon broader spaced resource drilling. The extraction process used a nearest neighbourhood algorithm, repeatedly moving the origin in 5 mE, or 5 mN increments. Individual resource estimates were then completed for each of the extracted drill hole data sets (49 in total) as well as a grade control estimate based upon the exhaustive data set. Whilst the data unpinning each of the 49 estimates changed, the geological assumptions, variography, and modelling parameters remained constant. This approach was taken to isolate the impact of changing the input data. The 49 estimates were then compared against each other and the grade control estimate. The mean of all 49 sensitivity estimates was close to that of the grade control estimate in terms of contained gold, tonnes and grade. Whilst it is acknowledged that the grade control estimate itself is subject to some degree of estimation uncertainty, the close match between the average of the 49 estimates and the grade control estimate suggests that the resource estimation methodology is reasonable and appropriate. Whilst this comparison is important, the focus of this study is on the component of estimation uncertainty related to the underlying data, and this is reflected in the spread across the estimates. The spread across the 49 global estimates (highest to lowest) for this particular Mineral Resource Estimation Conference 2023 | Perth, Australia | 24–25 May 2023 142 case study was found to be significant (approximately 20 per cent in grade and metal) and that is attributable solely to the underlying data. This exercise quantifies a component of the estimation uncertainty that is inherent to all drill hole data and is distinct from the uncertainties associated with modelling methodology choices, sample and subsample quality and drill hole spacing-related interpolation uncertainty. Importantly, this uncertainty is unknowable prior to mining; we can only directly compare the resource drill hole data against grade control data after mining has taken place. For the case study, approximately 65 per cent of the estimates fell within 5 per cent of the grade control estimate for contained-gold, suggesting that for many projects the histogram of drill hole data is unlikely to differ noticeably from that of the in-ground resource. However, about 15 per cent of the estimates in this study differed by more than 7.5 per cent, suggesting that a not-insignificant proportion of estimates will be materially compromised. Whether or not the available drill hole data is representative comes down to the ‘luck of the draw’ and cannot be known at the time of resource estimation. Given the challenge of attempting to evaluate forward-looking estimation uncertainty, a component of which that can only be quantified retrospectively, what should we do as resource geologists? What are the implications for risk evaluation, resource classification and reconciliation? Furthermore, without rigorous post-mining data checks, resource geologists may conflate suboptimal modelling with the shortcomings of the underlying data. An example of this problem is discussed in the following section.

May 24, 2023

Forty two new apatite and zircon fission track analyses on the Kubor-Marum-Bena Bena area of northern Papua New Guinea severely constrain the timing of tectonic events and hence the tectonic models for New Guinea, particularly arc-continent collision. In addition, the data can be analysed regionally in terms of known and assumed lithospheric fault zones inferred to control mineralisation (eg. Corbett 1994; Kendrick et al 1995; Hill et al 1996) to determine reactivation of these zones. When combined with the new kinematic plate models for the evolution of New Guinea (Hall 1997) there is the potential to predict likely zones of mineralisation. The low-medium grade? Palaeozoic rocks of the Bena Bena terrane yielded a zircon fission track age of 17 Ma, indicating the time of cooling below 200¦C. The age is consistent with `the shift, at about 16-18 Ma, from volcanolithic sediments to mixed provenance sediments rich in quartz and metasedimentary lithic fragments in the southern Finisterre Ranges (Abbott et al. 1994). This shift was recorded in the Sukurum Formation, which ranges in age from Middle Miocene to Pliocene with quartzose and metasedimentary grains in all samples. Although recording rocks at temperatures of 200¦C and hence depths of ---5km, the 17 Ma age suggests the Bena Dena terrane was rapidly cooling at that time due to uplift and erosion, supplying metasediments to the southern flank of the Finisterre Range, currently 100 km to the east. Significantly, the Middle Miocene Akuna Intrusive Complex, dated 14-17 Ma by Page (1976), lies almost entirely within or adjacent to the Bena Bena terrane. Potentially the uplift, erosion and cooling of the terrane at 17 Ma may have been caused by the rising plutons. If so the 800 sq km exposure Akuna Intrusive Complex at the SE end of the Bena Bena terrane may have been the focus of uplift and supplied the metasedimentary detritus to the Finisterre Range. This suggests that the Finisterre Range was not far removed from the Bena Bena terrane, perhaps less than the 200 - 400 km indicated by Abbott et al (1984). The Akuna Intrusive Complex probably resulted from subduction of the Solomon Sea Plate beneath the PNG margin. In order for the subducted slab to reach depths of -100km below the area to facilitate partial melting and plutonism (Hamilton 1994), subduction probably commenced 3-5 Ma Ma before plutonism, ie at or prior to 20 Ma suggesting a significant change in the tectonic regime at -20 Ma, perhaps related to collision of the Solomons - Ontong Java Plateau collision (eg. Hall 1997). The main result to emerge from the 28 apatite fission track analyses is that the whole of the study area underwent rapid cooling due to uplift and denudation in the Late Miocene, mainly between 7-10 Ma. Results from several samples indicate that they were not totally overprinted prior to uplift indicating maximum burial of -3 km in the Middle Miocene, assuming normal temperature gradients. The burial is less if higher gradients are assumed, allowing for Middle Miocene plutonism. The Late Miocene cooling in the Mobile Belt agrees well with the Pliocene to Recent cooling in the Fold Belt to the SW, interpreted to be due to uplift and erosion associated with fold and thrust deformation of the Miocene Limestones (Hill & Gleadow 1989). This timing relationship suggests, that compressional deformation in the Mobile Belt occurred in the Late Miocene, at 105 Ma, and that the deformation migrated towards the undeformed foreland in the southwest in the Pliocene, at 5-2 Ma. The regional Late Miocene uplift and denudation in the Mobile Belt is also consistent with that postulated by Crowhurst et al (1996; 1997) for northwest PNG, although they considered the event to have occurred mainly from 8-5 Ma. If so, then the uplift and denudation may have been migrating from east to west as well as to the south into the Fold Belt. These timing relationships place significant constraints on tectonic models for the area, indicating that the major compressional pulse in PNG commenced in the northeast at -10 Ma and migrated to the southwest and west in the latest Miocene to Pliocene. This suggests that the initial arc-continent collision and associated shortening occurred in the Late Miocene around -10-12 Ma. This immediately followed intrusion and cooling of the Akuna Intrusive Complex at -17-14 Ma and the Bismarck Intrusive Complex at -12.5 Ma (Page 1976), indicating a linkage between the end of plutonism and the start of compressional deformation. The timing is consistent with the observation of Hamilton (1994) that in areas of subduction `the common regime in overriding plates is extensional and leading edges are crumpled only in collisions. This suggests that the Middle Miocene magmatism in northern New Guinea occurred during extension above a subducting slab and was terminated by Late Miocene arc collision and resultant compression commencing at -10-12 Ma. In general, these data indicate that the acme of denudation and cooling in the Mobile Belt was in the Late Miocene, with relatively little denudation there since. In contrast, Hill & Gleadow (1989) found that the peak of denudation in basement outcrops in the Fold Belt was in the Pliocene at -4 Ma. However, in this study some samples from the Mobile Belt yielded Pliocene ages or high

Jan 1, 1997

Minerals and Metals Group (MMG) Century mine (Century) is a large open-cut mine in northwest Queensland approximately 2.5 km long, 1.5 km wide and 240 m deep, with an ultimate pit depth of 320 m. In 2009, the mine experienced three failures which began as small failures, and progressively developed into significant multi-batter failures. The incidents were managed by a combination of analysis, redesign and practical management to achieve positive outcomes for the mine. On the south wall of the mine, a series of batter-scale wedge failures occurred in December 2007. Analysis of structure orientations and design changes aimed to minimise the risk of failure extending over multiple batters, while accepting a risk of further small-scale failures. Design changes were constrained by a culturally sensitive area (CSA) located south of the wall. As mining progressed, shallow wedge failures developed over multiple batters and cracks were observed 60m back from the pit crest, indicating that a large-scale wedge failure was developing. Movement increased in January 2009 as a result of significant rainfall. After considering several design options, the area was stabilised by a cut-back of the wall. Two multi-batter failures also occurred on the north wall as mining progressed through a structurally complex area. Both north wall failures began as small failures that progressively developed, and required a redesign of the wall below as well as practical management of safety aspects. The complex structural geology was difficult to predict ahead of mining, and required geotechnical input into design changes as mining progressed. All three failures potentially had significant impact for ore recovery, and were managed by a combination of analysis, redesign and operational management have resulted to achieve positive outcomes in a challenging mining environment.

May 1, 2010

Historical exploration of the Ballarat East goldfield defined two primary quartz vein lode types. Tracking the intersection of these lodes with vertical lines of enrichment (or so-called æindicatorsÆ) was the basic practice for historical mining. This strategy proved highly successful in terms of identifying quartz veins associated with mineralisation. However, it provides little understanding of the timing or processes involved in the formation of these veins, or the associated emplacement of gold. Current underground mapping shows the Ballarat East gold deposit can be subdivided into four stages of vein formation. The earliest veins (V1 and V2) strike north-south and formed during bedding flexural slip and limb thrust fault propagation in parasitic chevron folds. V3 veins are subhorizontal and relate to further fold tightening involving a transition from east-west compression to north-south extension that initiated re-shear of bedded slip surfaces and additional strike-slip faults. Sporadic V4 veins overprint the shallow-dipping sections of V2 and V3 veins and are associated with weak north-south compression. The V4 veins mineralogy is distinguished by a high carbonate content and sphalerite-galena sulfide assemblage. Vein sampling indicates a significant increase in gold mineralisation where overprinting by the later deformation/veining has occurred. As a result of these observations, a new model is emerging that identifies the chronology, structural control and spatial distribution of respective quartz vein types. This is a major advance in terms of exploration targeting, resource modelling and mine planning.

Jan 1, 2008

By W Baguley

Attritioning has been a method of improving mineral extraction for over 60 years, however it is only recently that an understanding of the mechanism employed and the improvement in efficiency that have been achieved.The paper will discuss these improvements and how to overcome the problems that have been raised. At present there is a dearth of information on the correct design and operation of these machines.The paper will address these issues as well as fluid on the Zeta potential, micro shear, retention residence time and pilot plant operation.A study on the hydrodynamics of attritioner operation has been identified and will be discussed.Surface chemistry, clean lines of attrition material and the treatment of fines significant advances in the understanding of attritioning highlighted.CITATION:Baguley, W, 2013. Attrition scrubbers, in Proceedings MetPlant 2013 , pp 229-246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 15, 2013

By I Ametov, S Harmer, R Alford

A process that allows selective flotation separation of low Cu/S ratio minerals (chalcopyrite and pyrite) from high Cu/S ratio minerals (bornite and chalcocite) from a mixed copper mineral concentrate under controlled oxidative-reductive potential conditions has been developed. The process has been extensively tested on samples with varying mineralogy using laboratory batch flotation tests and mini pilot plant trials.In this process, the mixed final copper mineral concentrate is purged with nitrogen, and then treated with sodium dithionite to achieve a slurry Eh in the -200 to -100 mV standard hydrogen electrode (SHE) range. Flotation was then carried out using nitrogen gas. As a result, a low Cu/S ratio concentrate (<1.3) and a high Cu/S ratio (>1.6) tailing stream were produced. QEMSCAN analysis of the flotation products showed that more than 85 per cent of the chalcopyrite and 95 per cent of the pyrite reported to the froth concentrate. Chalcocite and bornite were almost evenly distributed between the concentrate and tailing streams.Surface analysis studies, focused on the mechanism of separation, showed that at a negative redox potential xanthate and possibly dixanthogen are the dominant hydrophobic species on the surface of chalcopyrite and pyrite. As a result, the flotation of these minerals remained unaffected. In contrast, a much lower concentration of xanthate species on the surface of chalcocite and bornite particles was observed. Therefore, it is likely that at low Eh the copper xanthate complex desorbs from the chalcocite and bornite surfaces, causing a decrease in surface hydrophobicity and the diminished floatability of chalcocite and bornite.CITATION:Ametov, I, Harmer, S, Grano, S and Alford, R, 2014. Selective flotation of chalcopyrite and pyrite from a mixed copper mineral concentrate under controlled redox potential conditions, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 131–140 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By U König

The use of high speed detectors made X-ray diffraction (XRD) an important tool for quality and process control in the mining industries. Besides knowledge about the mineralogy of a sample it provides useful information in terms of quantification of the crystalline phases and the amorphous content. Cluster analysis of XRD data can facilitate multi-dimensional mapping of ore deposits and drill cores, identifying regions of favourable mineral compositions. The studies of this paper demonstrate how XRD and data clustering can be used for grade control, process optimisation and quality control of iron ores and iron ore sinters. These new developments have an enormous potential as an inexpensive, reliable tool, useful in the characterisation of iron and sinter materials implemented in an industrial environment.

Jul 11, 2011