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By Dachun Liu, Baoqiang Xu, Bin Yang, Yong Deng, Heng Xiong, Xiumin Chen

"In this paper, vacuum distillation technology has been employed to study the possibility and regularity of separating copper and enriching lead, silver and other valuable metals from copper- based multi component alloys. The result of experiment indicated the practicability of vacuum distillation dealing with copper base complex multi-component alloy to separating lead and silver, when the vacuum distillation temperature was 1637K, distillation time was 60min; residual gas pressure was 20-30Pa, the removal rate of silver was 95.86%, of lead was 99.63%; the evaporation rate of copper was 0.745%. The removal rate of arsenic, antimony and tin respectively were 83.09%, 48.12% and 24.54%. Vacuum distillation method is applicable for dealing with the copper-based multi component alloys, silver and lead can realize effective recovery.IntroductionCopper generally go with lead mineral, therefore, to remove copper from lead becomes an inevitable link in lead smelting. Dross from removed copper usually contains lead, arsenic, antimony, tin and noble metal silver. In abroad, soda ash-iron dust is adopted to process copper dross of this kind. The outcomes of this process are lead matte, wet lead, scoria smoke dust etc[1~5] The dominant components of all the outcomes are copper and lead. Moreover, in the process of tin smelting, copper contained by-products are produced, such as, crude tin alloy, tin lead electrolytic anode slime and so on. Copper content of all the products are ranging from few percentages to several tens percentage. Those products, no matter merge into main producing process, or separate treatment cost losses of valuable metal. For one hand, the containing of copper of outcomes from pyrogenic attack is less than 50%, for the other hand, a lot of valuable metal need to be processed, which increase the cost of retrieve."

Jan 1, 2013

By J. Kent Smith

THE object of this investigation was to ascertain the effect of varying percentages of beryllium upon pure copper and the properties of the resultant alloys in their softest condition, the effect of heat hardening on them, and the extents to which these properties could be augmented by different combinations of heat hardening and cold rolling. The data compiled in this paper represent partial results of continuing work done by the writer during the years 1930 and 1931, which was a continuation of the unpublished ground work incepted in the laboratory of the Beryllium Corporation of America by H. B. Pulsifer in the year 1926. A short bibliography on beryllium alloys is appended to this paper. It has, of course, long been known that copper subjected to cold work is hardened and strengthened to a considerable extent, furthermore it had been qualitatively noted in previous experiments that copper-beryllium alloys seemed to take on hardness and strength as the result of cold work at a greater rate even than does plain copper. It was decided therefore to endeavor quantitatively to determine such rate of increase, as also the relative increases by combinations of cold working and heat hardening. The results are summarized in Tables 3, 4 and 5. A preliminary series of experiments had indicated that by an exposure to the heat of 575° F. for 2 1/2 hr. age-hardening of a 2.50 per cent beryllium-copper alloy took place. Slab ingots of nine different beryllium-copper alloys were prepared under identical conditions by means of the addition of a previously prepared 12.5 per cent beryllium-copper alloy to well melted and deoxidized pure copper. The ingots were then reduced by hot rolling to 12 gage (about 0.080 in.). These strips were heated to 1450° F. and quenched in cold water, and thus brought to their softest condition. (See Table 1.) Strips were next "precipitation-hardened" by reheating to a definite temperature. Similar strips to those itemized in Table 1 (which had been quenched from 1450° F.) were reheated for varying periods of time at 575° F. The hardness figures show that a metastable condition, of which the range of stability is dependent entirely upon time at the given temperature, begins to be noted in copper samples containing as little as 0.85 per cent beryllium. It is extremely noticeable in samples containing

Jan 1, 1932

By Zay Jeffries

AN acute current shortage of copper, with the prospect that conditions may become worse, indicated by Office of Production Management information. Present estimates of copper requirement for defense indicate that nearly all the available metal will be required in the future for this purpose. Severe dislocations are, expected, but the objective is to distribute the copper available for civilian use in such a manner as to minimize these dislocations. The advisory committee on metals and minerals has been asked for a suggestions which may be helpful in meeting this situation.\

Jan 1, 1941

By W. H. (Bill) Jay

The destruction of cyanide creates a significant impost on the financial returns of gold mines. These losses increase substantially when ores containing cyanide-soluble copper are treated. Failure to remove or destroy free or WAD cyanide before it is discharged into tailings dams has, on a number of occasions, led to significant fish, bird and animal deaths. Thus, the retention of cyanide within the leach circuit and the recovery or removal of stable metal cyanide complexes needs to become standard practice for all gold mining operations. Furthermore, reducing cyanide purchases will minimise the quantity of cyanide transported to the mine site thereby reducing the opportunity for spillage of this chemical. Many different physical and chemical methods have been investigated for the removal of cyanide from process solutions. All cyanide destruction methods create a financial impost on the economic viability of the gold mine. To reduce these costs, the recycle of cyanide and the economic recovery of the metals solubilised by cyanide is of prime importance. The alternative methods currently in use, or under consideration for the treatment of copper cyanide, entail high reagent costs, produce the recovered metal in a contaminated form and require the safe handling of hydrogen cyanide. These processes will be briefly discussed within this report Environmental and financial benefits are enhanced by recent laboratory and pilot plant studies conducted by Oretek Limited. The Oretek CPC process for the recovery of copper and the recycle of cyanide from copper cyanide-containing solutions does not require pH adjustment to the alkaline cyanide leach conditions. Copper ions are coordinated to a water-soluble and/or an organic soluble polychelating agent under alkaline conditions and the cyanide ions are displaced Membrane separation, or phase separation of the polychelated copper from the cyanide ions enables the cyanide ions to be recycled and the copper to be directly electrowon from the polychelating polymer. The polychelating polymer is then recycled. Ion exchange resins and solvent extraction systems which may be incorporated into the above process will also be discussed in detail.

Jan 1, 1998

By Wong PLM, Hsu YJ, Nguyen HH

The treatment of copper-gold ores has been plagued by several problems because of the interaction of copper minerals with the cyanide leach solution and the dissolved gold cyanide. Apart from the higher consumption of cyanide caused by the dissolution of various reactive copper minerals such as azurite, malachite, chalcocite, cuprite and bornite, native copper also causes the loss of gold due to its cementation reaction with the soluble gold cyanide. This interaction however is complicated by the strong dependence of various copper-cyanide species on the solution pH and free cyanide concentration.By understanding the effect of various operating conditions (pH and cyanide concentration), several treatment processes could be implemented to minimise the copper-gold interaction and to separate copper from mixed copper-gold cyanide solutions.

Jan 1, 1997

By LaBrooy S. R, Ritchie I. M

Problems with the cyanidation of copper-gold ores Copper-rich gold ores are becoming increasingly important both within Australia and overseas (Dayton, 1982; Deng and Muir, 1994; Muir et al, 1993; Sceresini and Staunton, 1991; Solis, Hefter and May, 1995; Wall, Hornby and Sethi, 1987). In Australasia alone over 30 copper-gold deposits have been reported (Muir et al, 1993). Table 1 lists several of the deposits that have been processed. TABLE 1 Examples of copper-gold deposits in Australasia. Name

Jan 1, 1997

By Steve R. La Brooy

Techniques for dealing with copper-gold ores are reviewed. Cyanide has little selectivity of gold over copper, and depending on the reactivity of the copper minerals cut-off grades for gold recovery by direct cyanidation are generally below 0.5% copper for oxide ores and 1 % for sulfide ores. Non-cyanide gold leachants also tend to be reactive to copper. The cheapest option for sulfidic ores is to produce a flotation concentrate for smelting, while transition ores generally need a sulfidisation pre-treatment. The requirement for the concentrate to have a high enough copper grade to be acceptable to a smelter may compromise copper and gold recovery. This can lead to significant gold grades in the flotation tailing, and copper levels in the tailing that are too high for direct cyanidation to be able to recover the gold. Oxidised ores, flotation tailing or other material not amenable to flotation represent the greatest grey area for gold recovery. Copper removal by acid, alkaline or ferric pre-treatment is possible though not generally used. Cyanide pre-leaching to remove copper, followed by copper and cyanide recovery represents a potential option that has been installed at one Australian gold plant, where copper is recovered onto carbon, ahead of the slower leaching gold. Actual cyanide recovery depends on ore mineralogy, allowing for losses as cyanate, thiocyanate, carbonate and ferrocyanide. Alternative options for gold recovery following cyanide leaching include precipitation, ion exchange and solvent extraction of the gold. Ammonia-cyanide leaching may offer another option for more selective gold dissolution from flotation tailing and oxidised materials. A plant is currently under construction in Mauritania.

Jan 1, 1992

By W. S. Fyfe, G. V. Albino, Ekawat Sitthithaworn

The mineralisation is associated with a composite diorite to quartz monzonite stock of Permian to Triassic age that intrudes Devonian carbonate strata. In the intrusive rocks it is associated with potassic alteration and in the carbonate wall rocks it is hosted by prograde garnet-rich and retrograde epidote-rich skarn. Both porphyry and skarn mineralisation include disseminated and fracture-controlled mineralisation, the latter probably accounting for the major part of the contained minerals. Copper, gold, silver and zinc are strongly positively correlated within the mineralised zones, although there are small differences in metal ratios between the different ore types; molybdenum appears to be negatively correlated with copper and gold. Ore minerals include chalcopyrite, magnetite and pyrite in porphyry-type mineralisation and bornite, chalcopyrite, pyrite, magnetite and lesser haematite in skarn.

Jan 12, 1993

By J. R. Long, T. R. Graham, R. S. Dean

ALTHOUGH considerable information has been published concerning manganese additions to the aluminum bronzes, these data refer principally to the two-phase alloys containing 8 to 11 pct aluminum, with some incidental information on the single-phase alpha solid solution alloys. Strauss1 gives an excellent review of the literature on aluminum bronzes up to 1927 and a bulletin2 of the Copper Development Association brings the bibliography up to 1938. The most extensive work on manganese addition is that of Rosenhain and Lantsberry,3 who investigated alloys ranging up to 10 pct aluminum and up to 10 pct manganese, and presented data on the physical properties of sand-cast, chill-cast, and several wrought alloys. The greater portion of the work, however, concerns alloys ranging from 7 to 10 pct aluminum in the cast condition and sheds little light on the properties of wrought single-phase alloys. Corson4 notes that alloys containing 10 pct manganese and 4 pct aluminum are not hardenable by heat-treatment, while one with 15 pct manganese and 5 pct aluminum can be hardened, but he gives no tensile data on it. In fact, no systematic study of the properties of single-phase wrought alloys containing substantial amounts of manganese has been reported. It should also be noted that most of the investigators worked with alloys containing significant amounts of iron and silicon, and that these impurities, whether intentionally added or accrued from the raw materials used in the preparation of the alloys, may be expected to have a direct bearing on the resultant properties. In the work on alloys made with electrolytic manganese, the Bureau of Mines has undertaken an investigation of the copper-manganese-aluminum system for the purpose of establishing the constitution and properties of high-purity alloys. In a recent paper5 the authors reported a metallographic investigation of the alpha solid solution area and the adjoining fields of the copper-manganese-aluminum system up to 50 pct manganese. This study established the alpha-phase boundaries at 50° temperature levels from 400° to 850°C. It was found that the addition of manganese to copper-aluminum alloys decreases the solubility of aluminum in the alpha phase (Fig I), curving the boundary toward the copper-manganese base line. The greatest change occurs in the first 20 to 25 pct manganese, which decreases the solubility of aluminum to approximately 4 pct. With larger amounts of manganese the alpha boundary tends to parallel the copper-manganese binary. At 1300°F and higher,

Jan 1, 1947

By J. R. Long, T. R. Graham, R. S. Dean

FOLLOWING the development of elec¬trolytic manganese production by the Bureau of Mines, an extensive program was planned to study the character of this high-purity product and its possible utilization in various alloy combinations. Previous work reported in the literature concerning alloys containing substantial amounts of manganese dealt largely with material prepared from relatively impure grades of the metal, and in many instances the impurities thus introduced did not permit true evaluation of manganese as an alloying element. One phase of the Bureau's program included a general review of the system and the establishment of the physical properties of copper-manganese-zinc alloys, using electrolytic manganese, and with particular emphasis directed to com¬positions within the alpha solid solution area. A number of reports of the work on alloys in this system, giving specific data concerning the equilibrium system and some of its alloys, have already been published. One study1 established the alpha solid solution area at various temperature levels and composition ranges up to 50 pct manganese. Other investigation reports presented the properties of selected alloys; namely, a 65 pct copper, 10 pct manganese, 25 pct zinc alloys' a sensibly white manganese brass composed of 70 pct copper, 20 pct manganese, 10 pct zinc,3 and a series of alloys with a constant copper content of 60 pct and 5 to 25 pct manganese, balance zinc.4 In these reports tensile properties as influenced by cold-working and subsequent annealing were given in detail. Additional data are now available on other alloys of the alpha region up to 50 pct manganese, and these, together with previously published information, are sufficiently complete to establish the relationship between properties and composition for this portion of the ternary system. The present report is a summary of the tensile properties of wrought alloys in the form of sheet and rod stock. The summary is necessarily limited to data for only one condition of cold-work and a single annealing temperature, since a complete exposition of all the data available would be too extensive for this presentation. Nevertheless, it is possible to make a reasonable approximation of the tensile properties in other conditions of cold-working and annealing by a correlation of the data previously presented on specific alloys and the data given here.

Jan 1, 1947

By P Thompson, D Rose, S Yu, W Baum, R Zahn, D Meadows, G Hall

Opportunities to improve mill operations are often cited by mining companies and recognised in many concentrators spread throughout the world but are not always followed through. Copper-molybdenum concentrator operations are challenged by a continuously increasing list of detrimental factors. They include deteriorating ore characteristics and poor mineralogical information on deeper ores, water quality and availability, steel wear and increasing power costs. Rapidly declining productivity, sustainability and environmental factors, new taxes, mining industry global skill set shortages, to name but a few, must also be considered. The days of maximising value by just increasing production or simple cost cutting are fading and will not be sufficient for the future of more complex mineral processing operations and larger plants. As a result, FLSmidth’s Global Process Technology Group has launched a plant and process optimisation program based on mineralogical and metallurgical surveys.Copper-molybdenum concentrators of the future need a paradigm shift. Modelling and simulation tools alone are not the answer to complex mineralogical associations and larger scale equipment and operations. Detailed area by area review aimed at identifying process constraints in terms of equipment and operating practice in conjunction with robust diagnostic work, ie mineralogical and metallurgical characterisation, are mandatory for process optimisation. The navigational beacon, for all operations who want to remain competitive, will be ‘process efficiency’. Since the use of automated mineral analysers has become more readily available and reliable, mineralogical-metallurgical plant surveys have proven themselves as the most efficient tool for optimising plant performance.The following paper will illustrate how operations can move toward eliminating under-performing flow sheets, fixing run-away flotation conditions, maximising individual process equipment performance, preventing shortcuts in Bond work index testing which may cause the failure of comminution models and minimising costly and time-consuming ‘debottlenecking’. As previous plant surveys have shown, the net project cash flow can be increased by 50 to 200 per cent if a combination of robust ore characterisation and plant process mineralogy is performed. The combined plant survey approach has been demonstrated successfully in a multitude of operations. This includes the use of key ore types in concert with statistically sampled process streams, bench-scale metallurgical tests and automated mineral analysis on a size-by-size basis.CITATION:Meadows, D, Hall, G, Rose, D, Thompson, P, Baum, W, Zahn, R and Yu, S, 2014. Copper-molybdenum concentrator surveys using mineralogy and process benchmarking tools to improve overall plant performance, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 315–322 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By Sadan Kelebek, Gerhard W. Heinrich, Peter F. Wells, Peter J. Whittaker

"Based on plant mineralogical data & circuit modelling, a pilot plant campaign was conducted at Strathcona Mill to identify options for improving the performance and operating flexibility of the copper-nickel separation circuits. Emphasis was given to options for combined treatment of the two feed streams (primary nickel ore rougher concentrate and copper ore circuit concentrate) to obtain more flexibility in the use of existing column flotation capacities. The pilot test results demonstrated that circuits which incorporated preconcentration prior to column flotation gave the best metallurgical results. Maximum column carrying rates were also related to the preconcentration. Predictions that increased circulating loads from the column scavengers would improve copper-nickel separation performance were confirmed.IntroductionThe copper-nickel separation circuits at Falconbridge's Strathcona Mill produce a high grade (>30% Cu) copper concentrate most of which is treated at the Falconbridge Kidd Creek smelter near Timmins. The remaining nickel concentrate which still contains approximately 20% of the feed copper is combined with other lower grade nickel/copper concentrates and shipped to the Falconbridge nickel smelter. The separation is required because of the limited capacity for copper processing in and downstream from the nickel smelter. Since no nickel is recovered at present in the Kidd smelter, the challenge is to minimize the nickel content of the copper concentrate.At present there are two independent separation circuits treating the rougher concentrates from the copper and nickel circuits as shown in Figure I. The minerals treated are mainly chalcopyrite and pentlandite but there is up to 25% of the copper present as cubanite in the copper ore. The circuits were kept separate because the flotation rates and froth characteristics were very different, probably because of the presence of cubanite and the large difference in the concentrate grades feeding the circuits."

Jan 1, 1995

By Leonard Larson

DURING several years immediately preceding the adoption of wet-charge smelting at McGill, various necessary conditions affecting this procedure, such as plant rearrangement and the metallurgical nature of the concentrate produced at the concentrator, were fully considered. Wet-charge smelting is not altogether novel to the recently reconstructed smelter at McGill. A number of years prior to this improvement, when flotation methods were initiated at the concentrator, large tonnages of wet and (at that time) sticky flotation concentrate were charged directly into the reverberatory furnaces, because of the difficulty in handling the entire tonnage in the roasters. The plant had excess reverberatory units and this surplus furnace capacity was utilized to smelt raw flotation concentrate. During the years 1927 and 1928, a number of actual operating furnace test runs were made to determine in principle the feasibility of wet-smelting operation. Results were encouraging, and in November 1932 the roasting plant was shut down. For some time thereafter the wet feed was delivered directly to the reverberatory side-charging calcine hoppers already in place. Feeding the wet material through the old calcine hoppers and charge holes into the furnace involved considerable labor, but under the conditions existing at that time it was desirable to continue wet smelting even under the handicap of this laborious procedure. The smelting plant during this period was on a curtailed production basis, operating only about 15 days per month. The alternate starting up and shutting down of the roasting plant presented a difficult problem and charging the wet material directly to the furnace entirely eliminated this difficulty. The roasting plant was never operated thereafter. EARLY PROCEDURE Fig. 1 shows the general layout of the smelting plant before the smelter was remodeled. Under this arrangement, all concentrates and fluxes except converter flux were delivered in cars to wooden storage bins, which paralleled the roaster building. The material to be roasted was dis-

Jan 1, 1938

By Leonard Larson

DURING several years immediately preceding the adoption of wet-charge smelting at McGill, various necessary conditions affecting this procedure, such as plant rearrangement and the metallurgical nature of the concentrate produced at the concentrator, were fully considered. Wet-charge smelting is not altogether novel to the recently recon-structed smelter at McGill. A number of years prior to this improve-ment, when flotation methods were initiated at the concentrator, large tonnages of wet and (at that time) sticky flotation concentrate were charged directly into the reverberatory furnaces, because of the difficulty in handling the entire tonnage in the roasters. The plant had excess reverberatory units and this surplus furnace capacity was utilized to smelt raw flotation concentrate. During the years 1927 and 1928, a number of actual operating furnace test runs were made to determine in principle the feasibility of wet-smelt-ing operation. Results were encouraging, and in November 1932 the roasting plant was shut clown. For some time thereafter the wet feed was delivered directly to the reverberatory side-charging calcine hoppers already in place. Feeding the wet material through the old calcine hoppers and charge holes into the furnace involved considerable labor, but under the conditions existing at that time it was desirable to continue wet smelting even under the handicap of this laborious procedure. The smelting plant during this period was on a curtailed production basis, operating only about 15 days per month. The alternate starting up and shutting clown of the roasting plant presented a difficult problem and charging the wet material directly to the furnace entirely eliminated this difficulty. The roasting plant was never operated thereafter.

Jan 1, 1938

By R. C. Wallace

Introduction At a time when the mineral-bearing areas in Manitoba are being fi very keenly explored because of the major developments which are now taking place in that Province, it may be timely to give an analysis of the mineralization in the two major types of deposits, with a view particularly to assist in the intensive prospecting to which the mineral-bearing areas in the Province are now to be subjected. As is already well known, there are two principal types of mineral deposits in the territory- copper-zinc and gold; and attention will be confined almost exclusively in this paper to these two types, with some indication of such modifications of the main types as may be found to be of economic value. The details of the paper are drawn from what is yet a small part of the area of the Province available for prospecting. Enough is now known of the types of mineral deposits which are described in the following pages to justify the expectation that these types may serve as criteria in the search in the newer districts. Doubtless new types will also be found as investigations proceed.

Jan 1, 1928

By Staunton W. P

The presence of cyanide soluble copper in gold ores can give rise to serious operational and economical difficulties in CIP and CIL operations. Cyanide consumption can result in the ore being uneconomic to treat by conventional means. Loading of copper onto carbon can also have deleterious effects on gold recovery and subsequent stripping and bullion recovery. The recovery of gold from high copper ores has been previously reviewed by Muir, LaBrooy and Cao (1989). Work undertaken for Mt Gibson Gold Project by Australian Mining Advisors with the assistance of the Chemistry Centre (WA) resulted in the patenting of a process and subsequent construction of a plant at Mt Gibson to overcome some of the difficulties encountered with copper. This paper discusses some of the work which was undertaken during laboratory investigations prior to commissioning of the plant and some further considerations for the treatment of high copper gold ores.

Jan 1, 1991

By C. E. Julihn

Technology concerns the ways of doing things; mineral technology the ways of performing operations required for obtaining minerals from the earth and extracting their valuable constituents for man's use. At any particular time, technology is relatively fixed, as everyone tries to adopt the way of doing things that has proved most profitable; but improvements evolve from minor changes here and there, and occasion- ally entirely new methods are discovered. Technologic improvement thus characterizes any long-time view of the mineral industries. Its test lies always in reduction of costs; there is no other incentive for the adoption of new methods. Reduction in cost of producing metal from ore decreases, in turn, the amount of metal that must be obtained from ore in order to pay for its production. An advancing technology thus is characterized by utilization of ores of lower and lower grade. A very important result of this is an increase in the gross quantity of metal obtainable at any price level, since the earth's crust contains vast amounts of very low-grade ores and relatively insignificant amounts of very high-grade ores. Technologic advances have, therefore, been the means of rendering the metals both cheap and abundant. Though concern as to possible exhaustion of the world's ore deposits is warranted, it is conceivable that a transcendent technology of the future might supply metal requirements from what we

Jan 1, 1932

By W. Warscheid

Coppernico fue evidenciado inicialmente como sistema tipo pórfido de cobre en el año 2008 por Quippu Exploraciones en virtud a un reconocimiento de cateos con mineralización oxidada de cobre en el valle de la quebrada de Choque, afluente del río Fortaleza. Dicha mineralización se desarrollaba en vetas de cuarzo hospedadas en granodiorita plutónica. El proEXPLO 2017 246 seguimiento de las mismas junto con el reconocimiento general del área permitió identificar un estilo de fracturamiento en enrejillado con relleno de cuarzo, óxidos de hierro y cobre en menor grado alojado en el cuerpo intrusivo.

May 1, 2008

By Yongfeng Jia

Lime neutralization and co-precipitation of As(V) with iron(III) is the most widely used and most economical method for the removal of arsenic from mineral processing solutions and effluents. However, the arsenate fixation mechanism in the coprecipitated solids is not fully understood. It was suggested by various researchers to be a ferrihydrite precipitation process during which arsenate is removed by adsorption on ferrihydrite. In this study, XRD, EXAFS and microprobe were employed for the speciation of arsenic and the effect of various factors on arsenic solubility was also investigated. The results indicate that coprecipitation is a more complicated process than simple adsorption resulting in lower arsenic solubilities. Poorly crystalline ferric arsenate was observed for the first time to be present in the coprecipitated solids. In case of lime as base, there exists some type of Ca-Fe-arsenate association. A general term ?coprecipitate? is suggested to describe the precipitated arsenic-bearing solids instead of those widely used specific terms ?arsenical ferrihydrite? or ?arsenic-bearing ferrihydrite?.

Jan 1, 2005

By G N. Musafer, M H. Thompson, R C. Wolff, E Kozan

The most important aspect of modelling a geological variable, such as metal grade, is the spatial correlation. Spatial correlation describes the relationship between realisations of a geological variable sampled at different locations. Any method for spatially modelling such a variable should be capable of accurately estimating the true spatial correlation. Conventional kriged models are the most commonly used in mining for estimating grade or other variables at unsampled locations, and these models use the variogram or covariance function to model the spatial correlations in the process of estimation. However, this usage assumes the relationships of the observations of the variable of interest at nearby locations are only influenced by the vector distance between the locations. This means that these models assume linear spatial correlation of grade. In reality, the relationship with an observation of grade at a nearby location may be influenced by both distance between the locations and the value of the observations (ie non-linear spatial correlation, such as may exist for variables of interest in geometallurgy). Hence this may lead to inaccurate estimation of the ore reserve if a kriged model is used for estimating grade of unsampled locations when non-linear spatial correlation is present. Copula-based methods, which are widely used in financial and actuarial modelling to quantify the non-linear dependence structures, may offer a solution. This method was introduced by Bßrdossy and Li (2008) to geostatistical modelling to quantify the non-linear spatial dependence structure in a groundwater quality measurement network. Their copula-based spatial modelling is applied in this research paper to estimate the grade of 3D blocks. Furthermore, real-world mining data is used to validate this model. These copula-based grade estimates are compared with the results of conventional ordinary and lognormal kriging to present the reliability of this method.CITATION:Musafer, G N, Thompson, M H, Kozan, E and Wolff, R C, 2013. Copula-based spatial modelling of geometallurgical variables, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 239-246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013