Search Documents

By M. Nowak

Extreme high grades often exert too much influence on surrounding block estimates, leading to a smearing of high grades that can cause overestimation of mineral resources. This concern underlies the decision to cap high grades, i.e., to lower the highest grades in a drill hole data set to some level deemed to be reasonable for producing reliable resource estimates. The choice of a capping level can, by itself, have a large influence on the analysis of the economic viability of a project. Despite this, it is often a choice that is made arbitrarily, guided more by art and tradition than by sound technical analysis. This paper makes suggestions for good capping practices, using methods and procedures that have been presented elsewhere, in the technical literature and in conference papers. The authors propose to use capping only if absolutely necessary, and review examples of cases where capping may not be necessary, or where capping does not make the estimates more reliable. It addresses two particular instances where capping is recommended: (1) where there is good reason to believe that estimation with uncapped grades will overstate the true underlying average grade, and (2) where there is potential for overly optimistic assessments of local block grades in the vicinity of some very high grade assays. The proposed capping practice is based on adjustments to only a very few high grade assays. The level of this capping can be assessed by bootstrapping and through analysis of the effect of extreme values on summary statistics like the average and the coefficient of variation. The paper recommends that good practice include an examination of the potential for local overestimation in higher grade areas.

Aug 1, 2013

Despite the claims of energy efficiency and the considerable development that has gone into dramatically extending rolls’ life and facilitating rolls’ change-out, in the six years since the last SAG conference, only a few HPGR operations have entered the realm of mineral processing. Risk barriers to the take-up of new equipment or processes are: a massive increase in materials handling issues as well as in the number of pieces of equipment required, low total circuit potential energy savings, and similar or higher capital costs. An alternative route is to make use of the complementary strengths of the HPGR and SAG in hybrid circuits. Minimising materials handling issues, providing an ideal balanced feed to the AG or SAG mill, and improving controllability through shifting grinding load as the RoM feed changes, are all potential major advantages to this approach. Conceptual circuit layouts, based on real measured ore and equipment data, are presented here to illustrate the potential production upside, the reduction in operating costs and the reduction in the physical and energy footprints.

Jan 1, 2012

By C. Jay Hodgson, P. J. MacGeehan, T. Sanders

"The orebodies of the Campbell and Dickenson mines are localized within an anomalously fissile and deformed zone that transects the eastern margin of a dominantly mafic volcanic complex close to its contact with a synchronously formed sedimentary basin containing clastic and chemical sedimentary rocks. The ore zones are mainly quartz-carbonate fissure veins and siliceous replacement zones formed during deformation and metamorphism, although there are some syn-volcanic exhalative ore zones in the mines. On a regional scale, the volcanic rocks have been affected by syn-volcanic spilitic alteration which has involved the leaching of Fe, Mg, Au and As from, and the addition of Na and variable amounts of Si to bleached zones enveloping primary permeable structures in the rocks. The leached constituents were deposited on the contemporaneous sea floor as iron-rich chemical sedimentary rocks. Although this alteration is not closely related, spatially, to the ore zones, it is characteristic of the general environment, and may serve to distinguish gold-prospective from gold-barren areas on a mining-camp scale.Closely associated spatially with gold deposits is a zone of feldspar-destructive alteration within which the volcanic rocks are characteristically feldspar-free, enriched in As, Au and to a variable degree in Si, and depleted in Na, Ca and to a variable degree in Mg, Fe and K. The alteration zone is also characterized by abundant, variably transposed, small quartz-dolomite to ferrodolomite veinlets formed during deformation and metamorphism, and includes, but extends outside of, the anomalously fissile zone hosting the orebodies. The alteration zone is interpreted to have formed syn-volcanically within the upper part of a major submarine hydrothermal discharge zone."

Jan 1, 1982

By O. Zach, K. Gamweger

"The Process of gas purging in molten Aluminium has been an industrial practise for many years. The melting of alloying elements and their rapid distribution in the bath is very important. To improve the mixing and homogenisation process, at RHI NON FERROUS METALS ENGENEERING GMBH a lot of CFD calculations were done. The so developed new ""purging program"" was tested at Alumino Catalan S.A., Spain. The results of the testtrials were very satisfactory and corresponds very well with the calculated data."

Jan 1, 2004

By Niels Verbaan, Ron Molnar, David Dreisinger

"An acidic chloride leach solution was generated from the treatment of an ore containing the following value elements: silver, gold, indium, gallium, copper, zinc and lead. Iron was also co-extracted into solution as a ferric chloride species. Iron extraction was a significant component of the overall acid consumption of the leaching process. The recovery of iron by solvent extraction and water stripping was investigated. Tri-butyl-phosphate (TBP) was very effective at recovering iron from the leachate. The water strip solution could be decomposed to produce ferric oxide (hematite) and hydrochloric acid for recycle via hydrothermal hydrolysis or pyrohydrolysis.INTRODUCTION The Malku Khota deposit is situated in the eastern part of the Bolivian Altiplano at elevations between approximately 3,800 and 4,580 meters above sea level (Figure 1). The project is located 98 km east-southeast of Oruro, and 85 km south of Cochabamba in a relatively remote area."

Jan 1, 2016

By B. S. Riess

A novel design for an automated uranium pellet stacking system is presented. This system is designed as a drop-in solution to the current production line to enhance the fuel pellet stacking process. The three main goals of this system are to reduce worker exposure to radiation to as low as reasonable achievable (ALARA), improve product quality, and increase productivity. The proposed system will reduce the potential for human error. This single automated system will replace the two existing pellet stacking stations while increasing the total output, eliminating pellet stacking as a bottleneck in the fuel bundle assembly process.

Aug 1, 2010

By N. Morrison, E. J. Rogans, A. J. Macintosh, N. Schoeman

"The design of carbon-in-pulp (CIP), and carbon-in-leach (CIL) adsorption circuits has essentially remained unchanged since the early days of the carbon process for gold recovery. For CIP circuit the pulp residence time is traditionally fixed at one hour per contactor with a carbon concentration of between 15 and 20 grams per litre, and the counter current carbon transfer rates are in the region of 15% pulp recycle. For CIL, the pulp residence time is usually three hours per contact or, with lower carbon concentrations ( 5 to 10 grams per litre).The drawbacks of using the counter current CIP/CIL adsorption circuits are discussed using modeling developed for CIP/CIL plant design. Counter current CIP/CIL, circuits are then compared with the carousel method of adsorption circuit design.The use of equipment suitable for the carousel approach is discussed which is, the AAC cylindrical and pumping (MPS) screens, the AAC pump-cell plant and launder designs.Finally, two plants at the Vaal Reefs Gold Mining and Exploration Company that use carousel adsorption circuits, namely the Vaal Reefs No. 1 Gold Plant, a CIL plant with stepped contactors, and the Vaal Reefs No. 2 Gold Plant, a CIP using a very short residence time (AAC pump-cell) are discussed."

Jan 1, 1998

By L. M. Dwarkin

Strip mining in Western Canada, though of considerable importance, is a fairly recent development for many of the operators in the district. Adding to the difficulty of unfamiliarity is the fact that a great deal of the stripping is done on pitching seams, for which the engineering and economic principles of setting the most advantageous strip ratio are not readily apparent. The novelty of the operation combined with the complexity of the principles involved can easily lead to misconceptions as to the determination of the economic ratio. Definition of Strip Ration There are two commonly accepted meanings of the term strip ratio: (1) The ratio of feet of overburden to be stripped per foot of coal recoverable. (2) The ratio of cubic yards of overburden per cubic yard or ton of coal. The first definition is inapplicable to pitching seams so, throughout this paper, strip ratio will be taken as the ratio of cubic yards of overburden to be removed per cubic yard of coal recoverable. Economic Ratio In the case of flat seams, the strip ratio is fixed by existing natural conditions, and the magnitude of this ratio determines whether stripping is economically feasible. With pitching seams, however, the strip ratio is a variable. On the outcrop, there is usually some coal which can be obtained with little or no stripping. As the work is carried down the pitch of the coal seam, the proportion of overburden continually increases until a point is reached where it is no longer economical to continue the operation. The question immediately arises as to the location of this point, and this resolves into the problem of determining the most favourable strip ratio, or, to use the accepted term, the economic ratio.

Jan 1, 1949

By Jozef S. Stachula

"INCO Limited operates eleven under-ground mines in the Sudbury area. The mines are located on the rim of the Sudbury Basin, an oval with the axis in the range of 27 km and 60 km. The ore dips to at least 3000 m below surface. The ores are mined primarily for nickel and copper. Total ore production from underground is in excess of 55 000tonnes per day.Over 40 surface fans have been installed since the late 1960s. All of the fans are adjustable pitch, axial flow units. A major factor influencing ventilation design in the last 30years has been the introduction of diesel equipment underground. Volumes per fan have ranged from 60 m3/sec. to 330 m3/sec., with motors from 100 hp to 2500 hp.Fans of the axial flow type have been commonly used for main fan installations at Canadian mines for many years. The standard arrangement has been to mount these fans horizontally, i.e. with the fan shaft and the long axis of the housing horizontal. This is a natural arrangement for an underground fan, but for a surface installation, a vertically mounted fan has definite advantages.The surface area taken up by a typical vertical fan installation is generally about one quarter that of a horizontal fan of the same capacity (Rutherford, 1957). This is not a problem with isolated fans and fat surface out crop sites, but where the installation is to be near existing buildings, or where there are poor surface soil conditions, space and cost considerations greatly flavour vertical fans."

Jan 1, 1996

By A. E. M. Warner, P. J. Mackey

Copper metal is produced in large or small tonnages in almost all countries of the world—handling raw materials in the form of either as-mined or milled ores, custom concentrates, or other custom materials including scrap and recycled copper. The scale of operations and type of technologies used around the world varies widely. There is a certain concentration of both mining and smelter-refinery production as a total of ten countries produce some 80% of world mined copper, while ten countries—mostly different ones— produce about 80% of smelted and refined copper. As a result, international copper trade is significant. The present paper takes the reader on a virtual tour of the copper plants throughout the world with a focus on smelter facilities. Operations and technology employed at each plant visited are discussed including a brief historical sketch. Future trends identified during the world tour are also discussed. The present authors considered sufficient time was required for a thorough study tour and settled on eighty days—the same as that for the celebrated fictional story of world circumnavigation in the 1870s. INTRODUCTION The celebrated French writer Jules Verne published his book “Around the world in 80 days” in English in1873 and it is the story of Phileas Fogg and his French valet Passepartout circumnavigating the world in 80 days in the fall of 1872 on a £20,000 wager (equivalent to over 2 million pounds today) from the exclusive Reform club in London. This is an interesting period for copper as it started trading on the London metals Exchange in 1876 and the start of global copper production chains started to take off at this time with the rapidly increasing electrification of the world. The production chain includes mining, mineral processing, smelting, refining and fabrication. Since 1880 annual world copper production has increased by over 150 times from 154,000 tonnes in 1880 (nominally considered as smelted copper—essentially the only route at the time to marketable copper) to over 23 million tonnes today (including total refined copper of 19 million tonnes of smelted copper including scrap and 4 million tonnes by solvent extraction-electrowinning), in tandem with increased electricity usage, along with significant transportation usage and for manufactured products (Figure 1).

Jan 1, 2019

By P. A. Riveros

Antimony is commonly found in copper electrolytes and can contaminate the cathode deposit. Several methods, such as ion exchange, solvent extraction and precipitation, have been developed to remove antimony from copper electrolytes. However, past research has been carried out almost exclusively with Sb(III), whereas the presence and behavior of Sb(V) in the electrorefining circuit have been largely ignored. In this study, the behavior of Sb(V) species was monitored and compared to that of Sb(III) species, using specially developed analytical and SEM-EDS techniques. Significant differences in the behavior of the two species were noted. Although the solubility of Sb(III) increases with temperature in synthetic copper electrolytes, the solubility of Sb(V) appears to be constant over the range from 30 to 80°C. Whereas Sb(III) forms several crystalline oxy-sulphate compounds, Sb(V) appears to remain as an amorphous oxide. Both Sb(III) and Sb(V) are extracted by aminophosphonic resins, and Sb(III) is readily eluted by HC1 or complexing agents. By contrast, concentrated HC1 elutes Sb(V) inefficiently and complexing agents, such as EDTA and tartaric acid, do not elute Sb(V) to any significant extent. This observation implies that in industrial practice, unless special measures are taken, Sb(V) will accumulate on the resin, eventually impeding the ion exchange process.

Jan 1, 2007

By R. E. Whitehead

"The earliest gold-mining operations in Ontario included the Cordova, Deloro, and Sophia mines north of Lake Ontario (see for example, Miller and Knight, 1914). The Cordova deposit, worked intermittently between 1891 and 1940, is reported to have yielded 22 774 oz of gold from 127 670 tons of ore averaging 0.19 oz/ton Au (Gordon et al., 1979). The property remained inactive for many years. In 1988, Gunnar Gold Inc. and Mill City Gold Inc. conducted a drilling program along the main Cordova vein. The present study is based on core from three drill holes through the Cordova gabbro (Fig. 1) and the carbonate alteration zones enveloping the quartz-carbonate veins. The relatively narrow width of the alteration zone and the availability of continuous sections of core provided an opportunity to conduct a compact mineralogical-chemical study of carbonate alteration and its effect on the redistribution of rare earth elements (REEs). Although the Cordova deposit is not of great economic importance, this study has revealed a number of interesting features which are interpreted as resulting from processes that may be applicable to gold-bearing quartz vein deposits elsewhere.The Cordova gabbro is layered in places, varying from fine-to coarse-grained and in color from light gray (feldspathic) to dark gray-green (amphibolitic). However, layering within the section of the gabbro considered here and which hosts the Cordova gold deposit is mainly textural rather than compositional. Samples of unaltered gabbro selected for analyses were medium- to coarse-grained; amphibolitic and anorthositic phases were avoided as being less representative of the general chemical composition of the gabbro.The Cordova intrusion has undergone regional metamorphism and consists dominantly of actinolite and saussuritized plagioclase with smaller amounts of granular clinozoisite and flaky chlorite; minor quartz, ilmenite, titanite, apatite, and calcite are typically present. Some actinolite grains have irregular rims of pleochroic green hornblende that is optically continuous with the actinolite. Both chlorite and clinozoisite occur in two modes. Chlorite occurs as an alteration product of actinolite and as separate grains in the groundmass. Clinozoisite and albite occur together as regional metamorphic alteration products of original plagioclase crystals and as discrete grains in the matrix.The main gold-bearing quartz-carbonate occurrence occupies an easterly trending, steeply southward-dipping carbonatized shear zone up to 15 m wide and 700 m long. The gold occurs principally in irregular lenses and veins of quartz-carbonate up to a foot thick occupying the lower part of the schistose shear zone within the gabbro. This zone consists of chlorite, biotite, and carbonate and is flanked on both sides by chlorite-carbonate schist lacking biotite. The chlorite-carbonate schist in turn grades into regionally metamorphosed actinolite-bearing undeformed gabbro.The dominant sulfide within the quartz-carbonate lenses and in the surrounding schist is pyrite which occurs mainly as small cubes (1 mm to 2 mm) generally making up less than 10% of the rock. Rare blebs of chalcopyrite are present. Magnetite is common throughout the schist."

Jan 1, 2001

By P. C. Sui

Numerical investigation and parametric simulations of the coupled electrical conduction and mass diffusion in the GDL of a PEMFC are reported. The coupled transport is solved using a 2D, "under-the-rib" computational domain for the cathode GDL. The current density on the GDL/catalyst layer interface, which constitutes one of the boundary conditions for the GDL domain and reflects the activation overpotential in the catalyst layer and the ohmic loss in the membrane, is solved iteratively using a novel numerical algorithm. A parametric study is performed to investigate the effects on current density distribution of various operating conditions such as oxygen concentration and membrane resistance, and of design factors such as GDL geometry, anisotropic transport properties, and deformation under the land area. The results show that the current density distribution under the land area can be dominated by either electron transport or mass transport, depending on the operating regime.

Jan 1, 2005

By W. B. Boggs

A paper describing the Noranda smelter was presented at the 1930 annual meeting of the Institute (1). A lacer article appeared in the Trans-actions of the A.I.M.&M.E. in 1933. Since that time there have been many changes and improvements in plant operation and it is the abject of this paper to give an account of developments up to the beginning of 1939. The Noranda smelter was built originally for a capacity of 30,000 tons a month. At present, the average tonnage being treated is 110,000 tons a month, the tonnage of custom ores alone amounting to almost the total rated capacity of the original plant. This increase has been accomplished by increasing the capacity of the reverberatory furnaces from 700 tons of solid charge per furnace day in 1928 to 1,500 tons per day at the present time. In 1935, it appeared chat the limit of furnace capacity had been reached at about 1,300 tons of charge per day because of the short life of the sprung silica arches. However, the adoption of the suspended magnesite roof, which had been developed at the Flin Flou smelter of the Hudson Bay Mining and Smelting Company, overcame this difficulty and it is believed that the limit of the capacity of the present reverberarory furnaces has not been reached. A brief account of the changes to equipment in the various departments will be given, followed by a description of the more important features of present smelter practice.

Jan 1, 1939

By K. E. Prince, S. Burling, R. J. Ring, M. K. Maley, W. R. Taylor

Vanadium occurs as a component of carnotite, which commonly occurs in calcrete type ores, and in association with uranium in some sandstone-hosted deposits. Increasing demand for vanadium, which is used in high-strength steel and Vanadium Redox Flow Batteries (VRFBs), is offering some uranium producers a potential option for by-product vanadium recovery to improve project economics at a time when uranium prices are low. This paper presents the results of recent optimisation studies on the Bigrlyi deposit for the recovery of both uranium and vanadium. The impact of carbonate leaching conditions on uranium and vanadium extraction from a range of carnotite-in-calcrete ores is also examined, with vanadium extraction limited by the proportion of non-carnotite vanadium present in the ores. The findings are particularly relevant to developing projects as results show that elevated vanadium content in ore does not necessarily equate to a high vanadium extraction in leach.

Jan 1, 2020

By Mariam Melashvili, Chris Fleming, Rene Jackman

"A study was undertaken to develop a flowsheet for treating a high grade silver/gold ore body by cyanidation, CCD and Merrill Crowe precipitation, with tailings detoxification to meet stringent limits for cyanide and thiocyanate concentrations in the final discharge. In one of the flowsheets examined, free cyanide in the final leach liquor was recycled to the mill where it was hoped that it might react beneficially with fresh ore to reduce the cyanide makeup by an equivalent amount. However, laboratory results suggested that most of the cyanide recycled to the mill reacted with iron forming a soluble ferrocyanide complex. This would have had a negative impact on cost, not only because the cyanide recycled to the mill could not be claimed as a credit, but also because the presence of ferrocyanide in the final tailings meant copper sulphate had to be added to the SO2/air cyanide detoxification process to precipitate the iron cyanide. In an alternative flowsheet that was examined, recycling of cyanide to the mill was abandoned, and this allowed the milled slurry to be pre-aerated in a single mixing tank before cyanidation. During alkaline pre-aeration, any ferrous ions that are generated in the mill are oxidized to ferric ions, which, unlike ferrous ions, are inert in cyanide solution. This significantly reduced cyanide consumption and lowered costs in the detoxification of the tailings. In this paper, the cost implication of these two circuits is compared.INTRODUCTIONMost of the gold that is produced by hydrometallurgical processes is recovered by cyanide leaching. The reason for that lies in the superior thermodynamic stability of the gold cyanide complex and the favorable economics this confers on the process (Fleming, 2014). However, the gold in most of the gold-bearing ores processed is a very minor component, and therefore, the leaching response is often dominated by that of the host minerals.Many gold ores treated by the cyanidation process contain iron minerals. Cyanide reacts readily with some iron minerals forming ferrocyanide as one of the compounds. Ferrocyanide is stable when solutions are exposed to air and does not retard the dissolution of gold. In fact, it has been claimed that potassium ferrocyanide in the presence of oxygen is a weak solvent for gold (Hedley & Tabachnick, 1958). The main problem with the formation of this compound is the consumption of cyanide. Ferrocyanide contains six molecules of cyanide for each atom of iron {Fe(CN)64-} so the wasteful consumption of cyanide is greatly increased by its presence. Thiocyanate is also often produced during the reaction of certain iron sulphide minerals with cyanide, and contributes to the wasteful consumption of cyanide. In addition, environmental regulations in certain countries demand that thiocyanate be removed from the tailings, and this adds significantly to the cost of tailings treatment."

Jan 1, 2016

By M. D. Lawton, P. Steen

"Drilling results from the established levels at AngloRouyn Mines Limited had indicated inadequate tonnages to meet the planning schedule. It was decided to develop the main zone of the orebody below the bottom working level of the mine. The orebody dips north at an average of 50 degrees and plunges to the west at about 10 degrees. Due to the need for rapid development of this area, a decision was made to use trackless equipment.This paper describes the method used and highlights the use of small 5-ton diesel trucks as carriers for the 2-cu.yd LHD machine. The method achieved advances of about 1,000 feet per month, at a dip of 12 degrees over a length of close to 2,000 feet, using as few as 20 men per day."

Jan 1, 1971

By C. R. van Staal, L. R. Fyffe, J. A. Winchester

"The geochemical characteristics of Early Paleozoic volcanic rocks are used to establish the following tectono-volcanic regimes in the northeastern mainland Appalachians: foreland continental rift volcanics; fore-arc ophiolites; island-arc volcanics; back-arc ophiolites; and back-arc ensialic volcanics. A Late Precambrian calc-alkaline magmatic arc bounds the Early Paleozoic volcanics to the southeast.The Precambrian calc-alkaline volcanics of the Caledonian Highlands, primitive Cambro-Ordovician oceanic-arc volcanics of the Eastern Townships, and within-plate Ordovician volcanics of the Miramichi Highlands can be related to a progressive change in mode of subduction from compressional Andean-type in the Late Proterozoic to extensional Western Pacific-type in the Early Paleozoic. The development of marginal oceanic basins in the Paleozoic created a more favourable environment for volcanogenic sulphide deposition, particularly in back-arc areas.The influence of tectonic setting is reflected in the distribution, size and metal ratios of the massive sulphide deposits associated with the volcanic rocks; deposits within intra-oceanic arcs in the Eastern Townships of Quebec average two million tonnes grading 2.2% Cu, 4.0% Zn, and 0.6% Pb, whereas those within a back-arc, ensialic rift near Bathurst, New Brunswick, average nine million tonnes grading 0.6% Cu, 5.4% Zn, and 2.2% Pb."

Jan 1, 1990

By N. Dave

"IntroductionIn the mineral processing industry, certain minerals such as uranium, thorium and vanadium, etc., are selectively extracted with high degree of purity from solutions containing variety of other impurities by solvent extraction process. This separation process consists of bringing two immiscible or partially soluble liquid phases in contact, where mass transfer of one or more components takes place by ion exchange under certain controlling parameters such as pH, salt concentration and temperature. The processes taking place are primarily physical because the solutes being transferred are ordinarily recovered without chemical change. For efficient extraction more than one stage is required and the solutions are recycled counter currently through other stages. A multi-stage counter current solvent extraction plants essentially consist of four circuits namely: extraction, water wash, stripping, and solvent scrubbing. In extraction circuit the desired mineral is extracted at a certain pH from an aqueous solution onto an organic solvent by bringing them in contact with mixing in large mixer-settler units. The mixed solutions are allowed to settle and separate into component organic and aqueous phases in horizontal settlers by providing sufficient residence time. The solutions are then moved counter currently to other units- the aqueous from the underflow and the solvent organics from the overflow of the settler. The solution streams coming out of extraction circuit are loaded organics and barren aqueous or raffinate solution. The loaded organics are washed with water in a water wash circuit to remove transported impurities from extraction circuit. In the stripping circuit, the desired mineral is stripped selectively from the solvent phase back into an aqueous phase with a stripping agent at a selected pH. The solution stream leaving, stripping circuit are stripped organics and pregnant strip (aqueous) solution. The stripped organics are scrubbed off other loaded impurities with a conditioning solution in, scrubbing circuit and the cleaned solvent is re-used in the extraction circuit.A typical solvent extraction plant for uranium extraction and purification at Agnew Lake Mine is described in this chapter."

Jan 1, 1989

By R. Mariano

"The ability to optimise a concentrator from mill feed to final products requires a process simulator that integrates comminution, classification and separation stages. A key part of any such integrated process simulator is the inclusion of a model of mineral liberation in comminution to predict the particle composition distribution of the comminution products. This need was recognised by Gaudin (1939) who proposed a model of liberation based on the cubic fracture of a mineral texture composed of cubic grains of two minerals.This paper describes the development of a texture-based model of mineral liberation that builds on the method proposed by Gaudin. The approach described here, termed the JK-Gaudin Random Liberation model (JK-GRLM) uses as its basis the measured, volumetric mineral texture data that can be quantified for a given ore using X-ray micro-tomography. The textural data required as inputs for the JK-GRLM include the mineral grain size distribution. While the Gaudin model (in common with other existing liberation models) is a two-mineral (or two-phase) system, the JK-GRLM extends the approach to multimineral systems of up to 15 individual minerals, making it more widely applicable.An example of the application of the JK-GRLM to a base metal sulphide ore is provided. In the example the JK-GRLM is used to simulate the breakage of the measured ore texture and the particle composition distributions of the simulated progeny particles are compared to the measured particle composition distributions of ore that has been physically broken.The ability to calibrate this mineral liberation model with measured ore texture will allow it to be applied in greenfield situations where geometallurgical programs measure texture and in brownfield applications as a key part of integrated process simulations. The development of this model enhances our ability to simulate mineral processing operations as an integrated system."

Jan 1, 2016