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By K Ringwood, J Cummings, C J. Moran

The development of a water accounting framework that enables consistent and contextual reporting of minerals operationsÆ water use, represents a key sustainable development objective that many Australian minerals companies have committed to. A future water accounting framework will require: key terms and definitions for water metrics for use in the Australian minerals industry that are consistent with the wider water sector, a suite of metrics representing key components of an operationÆs water balance, a calculation method for some of the metrics, and a reporting protocol. This paper presents the methodology that was used to derive these four components. Existing accounting frameworks were reviewed and key terms and definitions that were consistent with current regulatory requirements and voluntary approaches were proposed. A calculation methodology was developed to obtain some of the key indicators and variables that cannot be easily measured directly (eg water loss through evaporation). The adopted terms and definitions typically require a combination of empirical and estimated data. To test the proposed definitions and calculation methodology, values for water metrics were derived for eight sites, selected to test a variety of commodity and biophysical situations. The proposed framework and definitions could deal with the specific situations selected as case studies. The associated calculation methodology produced complete accounts at various levels. Some elements of the accounts could be directly checked against available data and simulations. The framework captured well variations in the reuse efficiencies; reuse efficiency could even vary for similar processes. It is concluded that this preliminary accounting framework provides a consistent language and metrics for quantifying and communicating water management, both at the intersection of the site with the surrounding landscape (through detailed reporting of site inputs and outputs) and within operational activities. Using this consistent framework for quantifying and documenting water use within an operation will provide a platform for compiling and comparing minerals industry water use with other sectors. This may support benchmarking activities within regions or companies and help demonstrate leadership amongst water users in water resource stewardship. The approach presented here is applicable to a wide variety of mining and processing activities, and biophysical environments, and following a period of stakeholder engagement and third party review, is being refined as a first step towards developing nationally consistent water use reporting for the Australian minerals industry.

Jan 1, 2009

By J L. Mauk

The Sovereign, Jubilee, Scotia and Jasper Creek prospects occur in an area of ~3.0 km by 2.5 km (~7.5 km2) in the Hauraki goldfield. The prospects are hosted by andesitic to dacitic flows, autoclastic breccias and localised pyroclastic and air fall deposits. Jubilee and Sovereign prospects lie along the northern and southern parts of the Waitekauri Fault, respectively, whereas Scotia is located ~1 km to the E, and Jasper Creek a further ~1.2 km to the ENE. Wall rocks are typically intensely altered, although the degree of alteration becomes more variable east of Scotia and at Jasper Creek. Alteration minerals include quartz, adularia, albite, chlorite, pyrite, illite, interstratified illite-smectite, smectite and calcite. Several of these minerals have distributions that show distinct zonation with adularia most abundant at Sovereign and Jasper Creek, but restricted to shallow levels at Scotia. Illite occurs throughout Sovereign and Jubilee plus the western margin of Scotia where it grades into interstratified illite-smectite; smectite predominates at Jasper Creek. Veins are uncommon and typically narrow, but show distinct zonations. Quartz veins are more abundant at Sovereign and Jubilee, whereas calcite veins are more abundant at Scotia and Jasper Creek. Laumontite veins occur at Scotia, and rare clinoptilolite, mordenite and stilbite veinlets with calcite occur at Jasper Creek. Fluid inclusion in quartz and calcite homogenised between 141¦ to 272¦C and trapped a dilute solution with an apparent salinity of less than 2.6 weight percent NaCl equivalent. Homogenisation temperatures are hottest at Sovereign (average 241¦C) and Jubilee (average 239¦C), cooler at Scotia (average 204¦C) and coolest at Jasper Creek (average 162¦C). Calculated paleowater table elevations suggest the water table was 680 m to 750 m above sea level at Sovereign and Jubilee, 450 m at Scotia and 225 m at Jasper Creek. The current data suggest three scenarios that can explain why Sovereign and Jubilee are deep and hot, whereas Jasper Creek is cool, shallow and marginal; 1.   some or all of the prospects may represent separate overlapping hydrothermal system, 2.   a high-relief hydrothermal system on the side of volcano with an inclined water table and outflow towards Jasper Creek, and 3.   a low-relief hydrothermal system that has been either tilted or block faulted with greatest uplift and erosion to the west.

Jan 1, 2005

By J Rutter, M NoppT

Ridgeway Gold Mine is part of Cadia Valley Operations which is wholly owned and managed by Newcrest Mining Limited. Ridgeway is an underground sublevel cave (SLC) operation mining approximately 5.6 Mt of gold-copper ore per annum. Ore tracking and accounting systems monitor gold and copper grades, copper mineralogy, host lithology and ore sizing. These factors have material impacts on concentrator throughput and metal recovery. Drawpoint recovered grades are routinely reconciled back to design data to monitor and assess draw performance. Grade control at Ridgeway has been based upon sampling the ore from SLC drawpoint bunds using 15 kg grab samples combined to form 30 kg samples prior to assay. This large sample size was selected in an attempt to representatively sample the coarsely-sized broken ore, the samples collected having a nominal diameter of about 130 mm. The assay results for gold and copper from these samples are extremely variable and the usefulness of the results has been questioned by the mine, particularly given the onerous and OH&S hazardous activity of handling large sample masses. Granulometry testwork at Ridgeway confirmed the extent of the relationship of decreasing sample grade variability with particle size. Based on this relationship, drawpoint sampling tests were carried out using 9 kg samples sieved to -45 mm at the drawpoint site. The gold and copper results from the finer 9 kg samples were significantly less variable than those from the normal, coarser 30 kg samples. The variance reduction was similar to that predicted from broken ore sample theory for the selected sample mass and nominal particle size. This relationship therefore allowed Ridgeway to assess the sample mass and particle size required to achieve a particular precision. In addition to improving the precision, the reliability of assay results from smaller samples from the fine particle fraction for drawpoint control depends on the accuracy of the results. The assumption is that the grade of the finer ore is similar to that expected from the lot, or that if a bias exists, then the extent of this bias is acceptable and consistent enough to be accounted for in the grade control system. The distribution of grade with particle size and the reconciliation of drawpoint sample and mine results is currently under investigation.

Jan 1, 2005

By M Rimes

This paper describes a technique to predict achievable head grades and likely variances from target grade for the Shovelanna Iron Ore Project, in the Pilbara region of Western Australia. The block model from which the production schedule is produced was created from a wide spaced drilling program and was classified as an Inferred Resource under the JORC Code. Ideally, production schedules and target grades would be determined using a combination of Proved and Probable reserves. However sometimes it is necessary to consider the requirements of potential off-take parties who have specific grade tolerances. A technique is needed to simulate mining sequences and grade variations early in the planning process, using an æexplorationÆ block model rather than the normal æminingÆ block model. Marketing contracts require examination whether the target grade {Fe, P, SiO2, Al2O3 can be kept constant, or must be revised during the life of the mine. A model (written in ExcelÖ using Visual Basic macros) develops a feasible, close to optimal, mine plan from block model data. Cut-off grades for minable ore are chosen. The model evaluates ore tonnage, grade and stripping ratio against depth. Outputs at this stage include plots of slices through the orebody, at nominated x, y and z coordinates, showing ore grade, waste, drill hole locations and topography. The model then searches for a feasible satisfactory mine plan. The initial grade target is the average for the identified ore blocks. The available block list (ABL) at any time is the set of ore blocks, any one of which can be mined without removing any other ore block. The entire initial ABL has an unacceptably large ætotal grade stressÆ, defined as [(Grade-Target)/Tolerance]2, summed over the four relevant minerals {Fe, P, SiO2, Al2O3}. The model trims the ABL, removing the block most harming the total stress. Trimming is repeated until the total grade stress reduces to an acceptable level. The trimmed ABL defines the first mine campaign. Removing the trimmed block set, a second ABL is exposed, which is again trimmed and mined. The process is automatic: a few seconds computation provides a complete mine plan. Re-running the model with adjusted grade limits quickly shows whether a uniform target can be achieved or whether the target must be modified during the mine life.

Jan 1, 2007

Solid EnergyÆs project evaluation process utilises a standard staged gate approach, with five levels leading ultimately to project execution (level 6). For a project to progress through each level, specific financial criteria need to be satisfied. Solid Energy primarily use value based (discounted cash flow) metrics to support investment decisions. Within this approach Solid Energy utilise a number risk-based techniques to quantify project uncertainty. This includes the use of risk-based probabilistic analysis using Monte Carlo simulation and probability-based scenario analysis. Figure 1 shows the net present value (NPV) cumulative probability. Project risks are identified through comprehensive risk workshops. The output of these workshops (list of key risks, ranges of uncertainty and probability of occurrence), is quantified within the financial models. Incorporating this risk information on a probabilistic basis allows an investment decision to be made with a robust knowledge of the range of possible project outcomes along with the likelihood (probability) of those outcomes occurring, for example 80 per cent confidence that the project value will be greater than NZ$10 M, 50 per cent confident it will be greater than NZ$50 M and ten per cent confident it will be greater than NZ$70 M. Incorporating risk implicitly into the financial analysis allows the business to easily evaluate/measure projects with different risk profiles. The detailed risk analysis does away with the complexity of arbitrary risk-based hurdle (discount) rates, as projects can be evaluated based on their specific risk characteristics. Whether a project will progress to the next level of investment and ultimately into implementation depends firstly on the value assessment in comparison with the associated risk/uncertainty, and secondly matching this with the businesses appetite for risk. At Solid Energy standard criteria are used for specific project evaluations, but in addition projects are considered on a portfolio or strategic basis also. Solid Energy attempts to maintain a mix of low risk (low value) projects with some high risk (high value) opportunities. The mix of projects must remain in line with the shareholdersÆ risk appetite. One of the keys to a successful investment decision is ensuring that the appropriate information is available to decision-makers so that they may proceed with a comprehensive understanding of the risks and impacts associated with the decision.

Jan 1, 2007

By B Ilich, M Brad, R Brandli, E MarT

The Wirtgen surface miner was investigated for application to the Cloud Break iron ore deposit since it can selectively mine thin seams and there is no requirement for drilling and blasting or primary crushing as the surface miner cuts and breaks the rock to secondary cone crusher feed size. A physical metallurgical study of the surface miner run-of-mine (ROM) from Cloud Break was conducted to better understand the product size distribution, amount of -31.5 +6.3 mm lump produced, lump stability and physical metallurgical quality compared to conventional drill and blast techniques. Since the Cloud Break mine was not in production, the lump product from crushing and screening of surface miner ROM was compared with product simulations using traditional premining physical metallurgical processing techniques including drop testing of PQ diamond core. Drop test conditioning is widely used in the iron ore industry to simulate blasting, crushing and screening at the mine and subsequent product degradation during handling from mine to port to predict both the amount of lump product on ship and the size distribution of the fines and lump. The size distributions of surface miner ROM for hard, medium and friable ore types show very tight size distributions with very little above 100 - 150 mm, about 60 per cent of the distribution between 75 and 6.3 mm, with less below 1 mm compared to conventional drill and blast ROM. Crushing, screening and drop test conditioning of the surface miner ROM produced more lump in equivalent ore types compared to drill and blast techniques or their breakage simulation using PQ diamond core. The higher lump yield from the surface miner was also confirmed by pilot scale crushing of some 7400 t of surface miner ROM from the Cloud Break trial pit. The lump produced from surface miner ROM was quite stable and showed similar low degradation characteristics to lump simulated from drill and blasting techniques. Product size distribution simulations using conditioned crushed lump from 1 m diameter Bauer hole samples were found to compare well with processed PQ core samples and surface miner product from trial mining at Cloud Break. The Bauer sample method is therefore considered to be a very useful close simulation to actual lump production, although the lump yield is lower.

Jan 1, 2007

It is difficult to evaluate exploration properties in a manner that is transparent, repeatable and easily re-calculated to cater for different geological interpretations, but the risk-based assessment methodology, the Geological Risk Method, developed and described by Lord et al (2001) attempts to satisfy all these criteria. This paper examines a case study example based on valuations of exploration projects conducted by SRK Consulting using this methodology, and describes some of the advantages and limitations of the method.

Jan 1, 2007

By Shillabeer J

Paste backfill has been adopted by many underground mines in North America and a growing number elsewhere. Paste offers a number of advantages. This paper suggests wider applications of paste technology to surface tailings disposal and water management. Advantages include improved overall stability, potentially easier permitting, smaller water inventory and earlier reclamation. It may be feasible to use the same plant to prepare backfill and to dispose of surplus tailings. Example cost models are given to illustrate potential applications.

Jan 1, 1998

By J Shillabeer

Paste backfill has been adopted by many underground mines in North America and a growing number elsewhere. Paste offers a number of advantages. This paper suggests wider applications of paste technology to surface tailings disposal and water management. Advantages include improved overall stability, potentially easier permitting, smaller water inventory and earlier reclamation. It may be feasible to use the same plant to prepare backfill and to dispose of surplus tailings. Example cost models are given to illustrate potential applications.

Jan 1, 1998

By T Wootton

The Barrytown ilmenite pilot plant was commissioned in two stages using Westport Titanium Project heavy mineral concentrate. The wet concentrator was commissioned on the 17 December 1989 and produced magnetic concentrate from this date.   The Drymill was commissioned on the 27 January 1990 and produced product grade ilmenite successfully from the onset.   The commissioning period was exceptionally short by normal standards and few problems were encountered. This paper concludes with general observations of operations of the pilot plant.

Jan 1, 1990

By T E. Hayward, D H. English, S J. Smith

Gold mineralisation at Bendigo is exclusively associated with quartz reefs, commonly found within or near anticlinal fold hinges. The gold is typically coarse grained (0.5 mm to 5 mm), free milling and displays a complex, nuggetty distribution within the veins. This ore is highly amenable to gravity recovery processes. As the Bendigo operation is not permitted to treat whole ore by a cyanide leach/CIL circuit, the flow sheet is more complex than conventional gold recovery circuits. Primary gold recovery is from a gravity circuit closing the grinding circuit to a bulk gravity concentrate and from a flotation circuit treating cyclone overflow to a bulk float concentrate. The plant capacity is relatively low by current gold industry standards at 600 000 tpa. Re-establishing gold production at Bendigo has presented several challenging, and sometimes competing, issues in process design and operation. The mine, process plant, associated services and infrastructure are at Kangaroo Flat, a low density, semi-rural zone in the southern urbanisation of the Bendigo metropolitan area. One of the main processing challenges during the design was the competing comminution requirement to avoid over-grinding, which would flatten or smear the characteristic nugget gold and hinder effective gravity gold recovery, whilst also providing sufficiently well liberated particles that would respond well to flotation. Important, non-metallurgical challenges were to minimise impacts on the local community from noise, dust, air emissions, visual impact, water egress and off-site traffic.

Jan 1, 2007

By Canterford J. H, Sparrow G. J, Dyson N. F

Pressure leaching of a bulk flotation concentrate prepared from a complex Cu-Pb-Zn sulphide ore was carried out at 150 and 190¦C under 1050 kPa oxygen overpressure in 1 L and 4 L titanium autoclaves to extract zinc, and to obtain residues for smelting tests and leach liquors for purification and recovery of the zinc by electrowinning. Dissolution of 95% of the zinc was achieved at both temperatures. Copper and iron dissolution and the degree of oxidation of the sulphide sulphur to elemental sulphur or sulphate depended on the temperature of the leach. At 190¦C, a jarosite (a basic iron sulphate) was the major iron-containing material in a residue that contained 17% total sulphur. Its formation was associated with dissolution of 18% of the iron. At 150¦C, oxidation of the sulphide sulphur and dissolution of the iron were significantly less. Only 7% of the iron was dissolved and a residue containing 25% total sulphur, present as a lead jarosite, lead sulphate, and elemental sulphur, was obtained. The formation of small pellets of unreacted sulphides coated with elemental sulphur caused problems during the leaching at 150¦C, and also during scale-up from the 1 L to the 4 L autoclave at 190¦C. At 150¦C the concentrate required two stages of leaching to achieve dissolution of over 90% of the zinc. These problems are discussed.

Jan 1, 1989

By G P. F Chitombo

The dilution of blasted ore with cemented backfill materials has been recognised as a significant factor influencing the economics of some underground metalliferous mining operations. Several mechanisms have been identified as possible causes of fill failures. The purpose of this study was to determine the possibility of damage as the result of one mechanism, that of production blasting adjacent to bodies of cemented hydraulic fill (CHF). A vibration amplitude transfer function has been developed to model the transmission of blast vibrations from the source, through ore, across the ore to CHF interface and into the CHF mass. The model may be applied in the near field. The model has permitted estimation of vibration amplitude at a point in a CHF mass, to which damage criteria may be applied. A number of assumptions and simplifications have been made to render the problem relatively tractable. Ore and CHF material characteristics recorded by a number of authors were utilised to formulate preliminary damage estimations. A simple application of the amplitude transfer model indicated that a zone of plastically deformed CHF would be produced, and would extend to a distance of approximately 0.5 metres from the interface into the CHF. A sensitivity analysis was also conducted, that indicated that the zone of plastic deformation would be more extensive given appropriate conditions. Recommendations for further model development and future laboratory and field investigations have been made.

Jan 1, 1998

By Chitombo GPF

The dilution of blasted ore with cemented backfill materials has been recognised as a significant factor influencing the economics of some underground metalliferous mining operations. Several mechanisms have heen identified as possible causes of fill failures. The purpose of this study was to determine the possibility of damage as the result of one mechanism, that of production blasting adjacent to bodies of cemented hydraulic fill (CHF). A vibration amplitude transfer function has been developed to model the transmission of blast vibrations from the source, through ore, across the ore to CHF interface and into the CHF mass. The model may be applied in the near field. The model has permitted estimation of vibration amplitude at a point in a CHF mass, to which damage criteria may be applied. A number of assumptions and simplifications have been made to render the problem relatively tractable. Ore and CHF material characteristics recorded by a number of authors were utilised to formulate preliminary damage estimations. A simple application of the amplitude transfer model indicated that a zone of plastically deformed CHF would be produced, and would extend to a distance of approximately 0.5 metres from the interface into the CHF. A sensitivity analysis was also conducted, that indicated that the zone of plastic deformation would be more extensive given appropriate conditions. Recommendations for further model development and future laboratory and field investigations have been made.

Jan 1, 1998

By A B. Christie, R Barker

Recent studies carried out to support the development of a strategy for the New Zealand mineral industry by the NZ Minerals Industry Association have included investigations of mineral potential, and the development of scenarios for realising this countryÆs mineral potential by 2010. New economic research is applying a general equilibrium model to quantify the opportunity cost of this potential. The model analyses separately four categories of land (private land and three categories of Crown land) and identifies four distinct mineral classes û precious metals, industrial minerals, coal and aggregate. It investigates the economic effect of the increased availability of mineral resources on the mining industry itself, on other industry sectors and on the NZ economy generally, including its impact on unemployment, GDP and household welfare.The analyses suggest that increases in the level of GDP of between 1.9 per cent and 3.4 per cent may be obtained by utilising New ZealandÆs mineral wealth to varying degrees. The mineral industry sectors are significant beneficiaries under all scenarios; mineral output increases by between two and five times its current level. However, the flow-on effects to the rest of the economy, and to households, are also significant. The analyses show that household income, adjusted for changes in consumer prices, can be expected to increase by between 1.6 per cent and 2.7 per cent. This results largely from increased employment.

Jan 1, 2002

By F Wu, H Wang, W Chen

A pressurisation method with multiparallel fans has been adopted at HongLiuLin Coalmine to control the significant external air leakage through the goaf area of coalface 15207. Potential damaging effects of large external air leakage on a coalmine’s ventilation system are analysed according to the current air leakage status caused by the exploitation of shallow-covered coal seams. The use of pressure balanced ventilation to prevent external air leakage through goaf areas is introduced. The calculation of ventilation network of HongLiuLin Coalmine shows that the air resistance of an external air leakage branch through a goaf area and the air pressure drops from the intake slope to the working face can be adopted to determine the values of pressure to be increased by regulating fans and airflow controlling parameters of regulators. By investigating airflow requirements of the pressure-balanced area as well as the performances of the auxiliary fans currently used in the coalmine, the number of pressurisation fans needed is fixed. The expected result is achieved after the pressurisation method has been carried out, which can be helpful for other coalmines to handle similar problems.CITATION:Wu, F, Chen, W and Wang, H, 2015. Multiparallel fan method to prevent significant external air leakages through a goaf area of a shallow covered coalmine, in Proceedings The Australian Mine Ventilation Conference, pp 267–270 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 31, 2015

By P Jonsén, S Larsson, M Parian, B I. Pålsson

The question of effective energy utilisation in grinding mills is not new. There are several conflicting arguments about tumbling mills, whether the efficiency is around 1 per cent or maybe 10 per cent, or even much lower. The energy not used is assumed to be lost as heating of the pulp, the grinding mill body, the charge, generation of shock waves and vibrations, etc. Stirred media mills on the other hand are generally considered to have better energy utilisation, but their energy efficiency is still not that clear. To shed some light on this a pilot-scale, wet stirred media mill was investigated over a range of operating conditions. The wet stirred media mill is a Drais PMH 5 TEX pearl mill fitted with an electric motor at 11 kW. It has been investigated over a range of operating conditions to try to balance the dissemination of the input energy in forms of the net grinding energy, mechanical energy losses, and the heating transferred to the pulp, the mill, the charge, and the cooling water. It is found that approximately 20–40 per cent of the input energy accounts for the grinding process. Also, that the difference between gross and net input electrical energy is mainly disseminated as heating of the pulp and cooling water. Mechanical energy losses appear to be much smaller than the heating effects. The use of a dispersant seems to mainly influence the heating effect.

Aug 24, 2022

By W N. Qin, B Zhang, W Z. Huang, X S Mo

Mengzi Mining and Metallurgy Co, Ltd (MMMC) polymetallic ore contains recoverable amounts of tin, lead, zinc, pyrite and other valuable minerals. With the change in orebody, the run-of-mine (ROM) ore changed from lead zinc rich ore to tin rich ore, where the current ROM ore contains 0.45–1.25 per cent tin, 20–25 per cent sulfur, and 0.20–0.35 per cent copper. The previous processing flow sheet was designed to recover lead zinc rich ore and it is clearly not suitable for recovery of the current tin rich ore, where the tin was already overground and tin recovery was only 50.19 per cent. The Mengzi Mining Company has recently modified the concentrator circuit by incorporating rod mill and high efficient Derrick fine screen. As a result of the improved classification efficiency, the overgrinding of the tin ore was minimised and the recovery of tin was improved by 20.71 per cent, ie from 50.19–70.90 per cent. An additional $10 M revenue per annum can be achieved as a result of the modification.CITATION:Qin, W N, Huang, W Z, Mo, X S, Zhang, B and Wheeler, J, 2014. Application of high efficient fine screen to improve the recovery of tin concentrate, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 487–492 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By J Li, A McFarlane, J Yu, L L. Kuhar, R Meakin, N Chapman

"Column leach tests are used frequently to predict recoveries and reagent consumption for nickel laterite heap leaches. These column leaches often run for more than 100 days and therefore require long-term monitoring and data analysis to achieve the desired information. To minimise the cost and time required to obtain relevant information for nickel laterite heap leaching, we have developed a small-scale hydrogeometallurgical laboratory test to serve as a proxy to column, and by inference, heap leaching. In excess of 50 leach tests were conducted using four different small-scale proxy and column leach tests to compare predicted nickel recoveries, acid consumption and leach kinetics and to determine whether insight into potential scale-up effects can be obtained. The most promising proxy test was identified and we achieved a statistically significant recovery and acid consumption correlations between this test and the column leach tests. The accuracy of the small-scale tests was confirmed by column tests run on new laterite ores. This geometallurgical proxy could be of use for the population of block models and process selection during the mine planning phase or for optimisation and control during processing.CITATION:Li, J, Kuhar, L L, Meakin, R, Chapman, N, McFarlane, A and Yu, J, 2016. A Geometallurgical Proxy for Nickel Laterite Heap Leaching, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 335–340 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

Australian rutile and zircon concentrates are refined raw materials and not manufactured products. The composition and properties of the grains are therefore fixed by nature. Specifications and assay certificates should take this into account by eliminating unnecessary provisions which only increase the time and cost of analysis.INTRODUCTIONThe raw material from which Australian rutile and zircon concentrates are isolated is beach sand on the coast of New South Wales and Southern Queensland. The principal current operations are in an area between Sydney and Brisbane.The sands consist principally of quartz particles, but they contain from 1 per cent to 50 per cent of a mixture of heavy minerals, including zircon, rutile, ilmenite, monazite, leucoxenic grains, chrome spinel, chromite, garnets, corundum, topaz, brown spinel, olivine, epidote, clinozoisite, tourmaline, magnetite, hematite, and cassiterite. The major components are zircon (25 to 60 per cent), rutile (25 to 50 per cent), and ilmenite (5 to 50 per cent).The so-called leucoxenic grains are not a definite mineral species but an alteration product of which the major components are titanium oxides and silica in varying proportions.Primary concentration is by gravity, using spirals or various forms of pinched sluice (Pullar, 1962), the resulting concentrates being redressed on wet tables. During these processes, some of the lighter heavy minerals including leucoxene and impure rutile, which tend to degrade the final rutile and zircon concentrates, are eliminated.The separation of heavy minerals from each other is...

Jan 1, 1963