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By Julie Hunt

Rimfire Minerals Corporation: early stage precious metal exploration specialists with discovery opportunities in partnership with major and mid-tier gold producers or prominent junior exploration firms. Mike Roberts, Senior Geologist

May 1, 2006

By C. V. Sibbald

"This paper provides a description of the Highmont copper-molybdenum concentrator from primary crushing through grinding, flotation, tailings disposal and water reclamation. The unique features of the operation will be emphasized.IntroductionThe Highmont property is located in the southern portion of the Highland Valley adjacent to and east of Lornex and about 4 miles south of Bethlehem Copper. The plantsite is at an elevation of 5100 ft.Highmont acquired the property in 1966 and in 1969 Teck Corporation entered into an exploration agreement which contained the option to finance the property to production . A detailed feasibility study of the project was completed in 1971. A positive production decision was deferred at that time because of the prevailing soft copper market and relatively low molybdenum price. The development of a strong molybdenum market in late 1978 and early 1979 led to a re-examination of project economics, and a decision to proceed to production was made in April 1979.The Highmont ore zones occur adjacent to and on either side of an east-west-trending quartz porphyry dyke. The two major zones , the east and west pits, contain 122 million tons of ore assaying 0.27070 Cu and 0.029% Mo. Chalcopyrite, bornite and molybdenite are the minerals of economic interest. Chalcopyrite is the predominant copper mineral. Mineralization occurs primarily along fractures and is zoned according to fracture density. Some pyrite is present, but for the most part it is outside the ore zone. Surface oxidation is shallow throughout the orebody.Wright Engineers Limited were responsible for the major part of the engineering design. Site construction, which began in the spring of 1979, is under the direction of Commonwealth Construction. The first concentrator line is scheduled for completion late in 1980, the other for early in 1981."

Jan 1, 1981

By D. M. Ward, J. H. Ahrens

"The Fond-du-Lac uranium deposit was delineated primarily by percussion drilling and a small program of diamond drilling. Drillcore samples from three diamond drill holes as well as digitized gamma logs from percussion and diamond drill holes were utilized to define and implement an algorithm for the conversion of gamma-log data to equivalent-grade data. The procedures and detailed analysis for the provision of assay data for uranium reserve estimation are documented.IntroductionThe Fond-du-Lac uranium deposit is on the south shore of Lake Athabasca, approximately 2 km southeast o f Fond-du-Lac , northern Saskatchewan. It is on the northern rim o f the Athabasca sandstone basin.A train of mineralized glacial erratics was discovered in 1967 by Farnok, Limited by means of a helicopter-borne radiometric survey. In 1969, a program of diamond drilling by Famok was successful in locating in-situ mineralization in the Athabasca sandstone at the apex of the mineralized boulder train. Since 1974, exploration and development work have been performed under a joint-venture agreement between Eldorado Nuclear Limited and Famok Limited, with Eldorado acting as operator.From the discovery drill hole to the development of the gamma-log to equivalent-grade con version algorithm, the deposit had been tested by a total of 222 drill holes, 210 of which were percussion holes and 12 of which were diamond drill holes. The holes are generally vertical and collared at 10-metre centres on fences spaced 20 metres apart. The deposit is approximately 390 metres long, 10 to 40 metres wide and up to 45 metres deep, and is generally confined to the Athabasca sandstone. Uranium occurs as pitchblende, coffinite and uraniferou s goethite."

Jan 1, 1983

By R. D. Studer

"The Snow Lake massive sulphide deposits are hosted by Aphebian volcanic rocks of the Flin Flon - Snow Lake greenstone belt, which in this area were deformed and metamorphosed to the amphibolite facies during the Hudsonian orogeny. All the known deposits occur in felsic rocks a short distance below the contact with an overlying mafic rock unit.The Stall Lake copper deposit occurs close to the top of an explosive felsic volcanic sequence. Rocks stratigraphically underlying the deposit have been hydrothermally altered by the mineralizing fluids. The sulphides were probably deposited from submarine volcanic vent(s) on the flanks of a felsic dome. The ore has been tightly folded, flattened and elongated parallel to the northerly trending lineation of the surrounding rocks during the several periods of deformation. Strike lengths of the lenses range from 60 to 300 feet and plunge-to-strike ratios are as much as 15:1.Ore lenses 1 to 4 typically consist of massive coarse-grained pyrrhotite, chalcopyrite and pyrite. A stratigraphically higher lens differs in that it contains fine-grained pyrite and chalcopyrite, with an increase in sphalerite toward its hanging wall. The No.4 orebody has a distinct footwall alteration zone characterized by brecciation and abundant chlorite-staurolite-garnet schist, with minor kyanite, anthophyllite, cordieriteand anhydrite, as well as sulphide minerals. Chemically, the alteration zone is distinguished from the unaltered felsic rocks by Na20 depletion and Mg-Fe enrichment.IntroductionThe Stall Lake copper deposit is located aproximately 4 miles southeast of Snow Lake in northern Manitoba (Fig. 1). The Snow Lake district has been a producer of gold (Rex, Bingo, Laguana and Nor-Acme Mines) and, since the fifties, of copper and zinc from deposits of Hudson Bay Mining & Smelting Co., Ltd . The district is located near the eastern limit of the Flin Flon - Snow Lake greenstone belt, which contains Amisk metavolcanic rocks of Aphebian age."

Jan 1, 1982

By M. J. Davie, D. S. Ross

"The successful design of a new mine water pumping system is in part dependent upon the accuracy with which the operating conditions can be forecast. This paper outlines the data which should be obtained on water quantity and quality. Methods of modifying the quality are also discussed. Some pumping systems for disposal of sludge are described, citing operating examples. The various types of mine dewatering pumps are described in terms of design characteristics as they affect operating efficiency and installation and maintenance costs. Piping system parameters are outlined. Finally, a hypothetical case study of pumping from a depth of 3,000 ft is assessed, with particular reference to capital costs.IntroductionMany mine operators would agree that the pumping system usually receives a low engineering priority during the design stage, and as a consequence causes some of the most frustrating problems during production. For a ""dry"" mine (up to 200 gpm), design of the pumping system may indeed be simple, but outside of the potash industry such mines are not common. For ""wet"" mines, the pumping system is important because it significantly affects the initial capital and on-going energy costs of the mine and, in the area of sludge disposal, the working conditions and production costs. Indeed, a failure in the pumping system usually has serious, and sometimes disastrous, consequences.The intent of this paper is to outline the factors which should be considered in the design of a pumping system and to describe the types of pumps used in mine dewatering and relate them to practical operating systems."

Jan 1, 1980

By G. P. Demopoulos

"The recovery and separation of platinum-group metals (PGM) from primary and secondary materials currently attracts a lot of interest due to the high value of the metals and their increasing demand for high-tech applications. Thus, the traditional refining processes are gradually replaced by newer more efficient separation techniques such as solvent extraction and ion exchange. In this paper the classical and modern refining techniques are reviewed and the pertinent solution chemistry dictating proper processing strategies in each case is discussed. The technological advantages of the modern separation methods are emphasized and reference to the recovery of PGM from primary as well as secondary materials is given. INTRODUCTIONDue to their outstanding physicochemical properties, such as catalytic activity, conductivity and corrosion/oxidation resistance, precious metals now find extensive usage outside of the traditional jewellery and investment industries. The increasing use of precious metals in such areas as electronics and catalysis, as well as their increasing high value, has led to a growing interest in their recovery and separation from primary and secondary materials.Quantitatively the demand of platinum metals in the western world by area of application is described in Table I. Platinum and palladium represent 90% of the total demand, Rh and Ru 8% and the remaining is taken by Ir and Os. In terms of value the following were the average prices (US$/oz) of the metals in the second half of 1986: 500 Pt, 125 Pd, 1200 Rh, 650 Os, 400 Ir and 80 Ru. Due to these prices, exploration, mining, reclamation and refining activity were very high in South Africa, the biggest producer of Pt metals in the world, as well as in other areas: United States (Stilwater deposit), Canada (Sudbury, Marathon and Lac des Isles, Ontario; Kenly Lake, Quebec, etc.), Australia, Europe and elsewhere"

Jan 1, 1989

By Min-Yuan Cheng

Heavy rainfall and typhoon oftentimes cause the collapse of hillslopes across mountain roads. Disastrous consequences of slope collapses necessitate the approach for predicting their occurrences. In practice, slope collapse prediction can be formulated as a deterministic classification problem with two class labels, namely ?collapse? and ?non-collapse?. Nevertheless, due to the criticality and the uncertainty of the problem, evaluating the collapse susceptibility of an area is a challenging task. This study proposes a novel Artificial Intelligence (AI) approach, named as K-Nearest Neighbor Based Bayesian Classifier (KNNBC), to deal with slope collapse assessment. In the proposed model, Bayesian inference is used as a framework to achieve probabilistic prediction of slope collapse. Meanwhile, K-Nearest Neighbor (K-NN) is employed as a density estimation technique. Equipped with probabilistic outputs, the K-NNBC is able to yield predictions with different levels of confidence and diminish misclassified cases. Experimental results point out that the proposed model is very helpful for decision-makers in slope collapse assessment and disaster prevention planning.

Aug 1, 2013

By H. D. S. Miller

"Artificial stope support has become an essential component of many underground mining operations. Because the use of cable bolts in open stopes is a relatively new concept, most support is still designed by trial and error or by rules of thumb. This often results in costly overdesigned support or inadequate support which leads to stope failure.This paper proposes guidelines to determine support requirements using the stability graph method. The stability graph method is an empirical open stope design technique that takes into account the key factors that influence stope stability. The factors for the design of a cable bolt support system include bolt length, bolt intensity and the original rock mass competency and can be estimated using empirically calibrated charts. The empirical charts were developed based on more than 60 case histories of cable bolting success and failure in Canadian open Slope mines.The concept of pre reinforcement and supp ort stiffness will also be discussed.INTRODUCTIONArtificial support, in the form of rock anchors, has become an important component of all underground mining operations because its stabilizing effect contributes to make underground workings safer. There are several types of rock anchors featuring different properties and having a variety of functions. This paper focusses on grouted cable bolt applications in large open stope mining. Such a support system may improve the competency of the disturbed rock mass to a point approaching the undisturbed rock quality by limiting joint movement and dilation. This results in more stable, possibly larger and, thus, more efficient production stopes."

Jan 1, 1989

By Lyle F. Smith

With the demand for larger diameter raises under all ground conditions and the increasing price of drill rods and operating costs, it was evident that there was a need for a raise borer with a variable-speed drive and fast-response torque limiting control. The raise boring machine described in this paper was developed after several years of market research and discussions with mine operators and construction contractors. This paper discusses the Ingersoll-Rand raise borers with the solid-state controlled DC motor drives. These machines have gained wide acceptance in the five-year period since being placed in operation and are designed to accommodate the range of raise diameters and lengths most required today.

Jan 1, 1978

By Paul Ransom

Sullivan Production and Present Value ? Mined 150 million tonnes ? 6.0% Pb, 5.5% Zn, 2 opt Ag, + Sn, Cd, In ? May 2006 Value + US$43 Billion ? 50% of metal from first 37% (55MT) of tons ? True original size ~ 110-120 Mt @ 17%Pb+Zn ? 30MT unconsolidated fill + HW cave

May 1, 2006

By G. M. Ritcey, M. Silver

"The processing of uranium ores produces tailings that could constitute environmental problems. At most operations the tailings are dumped into a designated area, with the ultimate acid drainage occurring if sulphides are present, together with the release of metal constituents to the rivers and streams. If the chemical reactions occurring within the tailings could be predicted, perhaps methods could be developed to either increase or decrease 'the rates of reaction. One such study has been in progress at CANMET during the past three years. The paper describes the use of Iysimeters containing low-grade and complex uranium tailings to determine the effects and interactions of bacteria, sulphides and solvent extraction organics on the release of radionuclides and on acid production. A simulated rain cycle nine-fold the normal was used and the Iysimeters were subjected to timed intervals of light and darkness. As a result, the tests simulated a 25-year period.IntroductionAlthough Canadians derive major benefits from uranium, there are environmental and health consequences, not yet fully understood, from the associated mining and milling operations. A major project at CANMET (Canada Centre for Mineral and Energ y Technology) has the ultimate goal of ensuring that any adverse effects on man and environment, now or in the future, resulting from the processing of uranium should be acceptably low. The magnitude of the task was exacerbated by the fact that the increased demand for electrical energy and the resultant increased prices for uranium oxide during the 1970s led to a rapid expansion in the industry. Annual world demand for uranium was expected to increase 12-15 fold in the next quarter century, with annual requirements of between 200,000 and 300,000 metric tonnes anticipated by the year 2000. Currently, the demand is low and the prices for the oxide have decreased."

Jan 1, 1982

By Wenbin Zhang, George W. Poling

"Prior or simultaneous addition of ammonium sulphate can, not only enhance the activating effect of hydrosulphide ions on xanthate flotation of malachite but it can also eliminate possible detrimental or depressing effects of excess residual hydrosulphide . Use of relatively inexpensive ammonium sulphate appears to reduce the consumption of the much more expensive sulphidizing agent for optimal flotation recovery INTRODUCTIONOxidized copper minerals float poorly with conventional thiol-type collectors. This is believed to be due to the nonadherent character of the collector coating, which is readily re- moved by particle abrasion. Rosas and Polingr reported emulsion flotation seemed capable of resisting abrasion during the flotation of non-sulphide copper ores. It appears that the most common flotation technique for oxidized copper minerals comprises sulphidization with Na .S and subsequent treatment by sulphydryl collector. The principles of this method are available . Malghan recently reviewed the studies on the role of sodium sulphide in the flotation of oxidized copper, lead and zinc minerals.It is well known that in the sulphidized flotation of malachite, the activation process depends strongly on the concentration of the sulphidizing agent, i.e., insufficient sulphide gives poor recoveries because of inadequate sulphidization; excessive sulphide also causes poor recoveries due to the depressive effect of the residual hydro-sulphide ions in solution. A major difficulty in sulphidization, especially with a varying feed nature, is to establish and maintain the optimum sulphide concentration in solution. In plant practice, vigorous control of sodium sulphide is required and stepwise addition of the reagent is recommended to improve the efficiency of the process. Soto and Laskowskim have found that malachite, depressed by an excess of sulphide ions, can restore its floatability if air or oxygen is introduced to the pulp. They attributed this to the oxidation of the HS - ions on the surface of the mineral. Jones and Woodcockuvi showed that a sulphide ion selective electrode could be used to optimize the conditions (e.g . amount of hydrosulphide, sulphidization time, flotation time) more readily and precisely than existing methods in the flotation of oxidized copper ores."

Jan 1, 1989

By W. J. D. Stone

Discussions with people knowledgeable in the insulating material field revealed that materials on the market would be considerably improved if there was a greater meeting of certain very desirable criteria. Such criteria include: (1) the material should not burn, even at higher temperatures; (2) at an elevated temperature, the material should not give off toxic gases or produce smoke; (3) the material should have a low energy input; and (4) vermin would find the material unattractive. Theoretical discussions led to the conclusion that an insulating material made by foaming a mineral, preferably of a platy nature, with an inorganic binding system that could be converted to a rigid structure at low temperature would likely meet all the above requirements. It was also theorized that the use of microwaves in the converting stage would be beneficial in the creation of a cellular structure by aiding the foamer through creating vacules of steam. Talc was selected as readily meeting the theoretical mineral requirement and the material of Baker Talc Limited was used as the basic filler. The early results obtained led to an agreement between Americ Mines Limited and Baker Talc Limited pertaining to commercial production in a specific geographic area. It was recognized that due to the economics of shipping, the availability of the basic filler material should be geographically widely spread. Samples of materials were requested and many received from all over North America and one from Britain. Enough work has been done to show that they respond to the basic process in a similar manner as the Baker Talc material. From the data acquired, it is considered that the process is applicable for the production of loose fill insulation, moulded insulation for pipes and structural steel, ceiling tile, batts, sheet and for incorporation into other products such as various roofing board materials. Sufficient development has been done to enable a budget to be prepared for the construction and operation of a loose fill insulation plant.

Jan 1, 1979

By James Logan

BC Alkaline CuBC Cu--Au PorphyriesAu Porphyries ?Formed over narrow time interval ~205 Ma ?from silica deficientvsalkali-enriched magmas ?At shallowdepths ?Structuralcontrol ?Cu-Au-Ag, trace Mo ?Oxidized system anhydrite, magnetite ?Multipleintrusions & multiplemineralized centres

May 1, 2006

By L. A. Clark, G. H. R. Burrill

"Most unconformity-type uranium deposits in Saskatchewan occur within a few tens of metres above and/or below the basal unconformity of the 1,45-b.y. Athabasca Sandstone. Graphitic basement rocks coincident with post-Athabasca faulting or brecciation at or near the unconformity are important in localizing uranium deposits which form as tabular, ribbon-like bodies with grades averaging over 2 per cent uranium and containing up to 50,000 tonnes UJOs. Some of these deposits have similar contents of nickel and arsenic.In the genetic model proposed to explain these deposits, traces of uranium were leached from the sandstone and weathered basement rocks by oxidized formation waters sweeping through the sandstone and along brecciated faults. Over basement meta-petites, a thick clay regolith adsorbed uranium from the solution. Thus fixed, the uranium was reduced through an indirect reaction with graphite, possibly involving methane. The clay mineral surfaces were thus continuously cleared to allow further adsorption. Basement faults localized along weak graphitic units were re-activated, offsetting the sandstone formation, probably at the time of an early uranium deposition about 1.2 b.y. ago, and served as vents for spent ore solutions. Fluid convection was induced by topographic relief and/or crustal heating from radioactive decay in basement plutons or mantled Archean domes with uranium and/or thorium contents a few ppm above those of adjacent rocks. Heat from radioactive decay within the incipient uranium deposit would develop in about a million years. The self-generating convection cell would continue uranium deposition until all permeability was plugged by minerals. In some instances, convection was re-initiated after subsequent faulting and brecciation. While some deposits were further enriched, heavy flows of oxidized fluids re-dissolved and partially depleted other uranium deposits.The East Alligator Rivers uranium deposits in Northern Territory, Australia, occur within Middle Proterozoic quartzchlorite and quartz-muscovite schists overlain by sandstone. Highest grades occur in silicified breccias where carbonate beds were leached out. Mineralization ages are both pre- and post-Kombolgie Sandstone, but, to date, no significant uranium mineralization has been encountered in the sandstone. There are many general similarities with Saskatchewan deposits, but important differences in detail."

Jan 1, 1981

By Naoki Takada, Dae-Sung Cheon, Minoru Sato, Masaji Kato, Manabu Takahashi, Weiren Lin

"The permeability and specific storage of rocks and sedimentary layers deep underground are critical parameters for issues related to the buried evaluation of natural gas and oil, CO2 aquifer storage, various kind of waste storage and so on.The Shirahama sandstone of the Wakayama prefecture, Kii peninsular in central Japan, consists mainly of quartz grains with no cracks and rock fragments that are aggregations of various kind of minerals. The quartz grains have an average size of about 150 micron. The pre-existing pore sizes were distributed dominantly at approximately 1 micron, and the total porosity was about 13%.During the hydrostatic compression test, we measured permeability of Shirahama sandstone by Transient Pulse test for three mutually perpendicular directions. Permeability of Shirahama sandstone remained fairly constant while increasing the effective confining pressure. Permeability anisotropy was observed slightly in the normal and two parallel directions to the bedding planes.We used micro focus X-ray CT with high resolution of 10 micron to visualize and measure the voids space information in Shirahama sandstone. Three-dimensional data were obtained in the intact and stressed sandstone simultaneously. Pore structures and grain boundaries were clearly distinguished. Pore closure and grain movement were clearly observed with increasing confining pressure. Micro focus X-ray CT is an effective tool to visualize and measure the geometrical information in rocks.Three-dimensional medial axis (3DMA) method was introduced to quantify the flow-relevant geometric properties of the voids structure in Shirahama sandstone using µfocus X-Ray CT data. The number of connecting path between two faces, tortuosity and the shortest path distribution within an arbitrary region of Shirahama sandstone specimen were analyzed. Geometrical information on the number of connecting path shows reasonable correlation between permeability anisotropy and mutually perpendicular directions normal and parallel to bedding planes. In addition, we introduced numerical simulation of fluid flow of pressurized Shirahama sandstone by LBM, to discuss the permeability change caused by pressurization on effective confining pressure. We obtained the detailed distributions of fluid pressure and fluid velocity in relatively narrow voids space more than 10 µm. It was confirmed that pressurization effect on permeability caused by the decrease of the connecting path and the shortest path between arbitrary faces, and then caused a complex condition on fluid pressure and fluid velocity."

Jan 1, 2015

By S. H. Lucas

"This paper describes the development and application of single-stage fluidized bed continuous ion exchange to the following processes: removal of base metals from mine water; cyanide removal from gold mill waste; uranium removal from sulphuric acid leach liquor; and uranium removal from mine water. IntroductionThe recovery of metals by fixed bed ion exchange has been practiced for many years. However, the operation has been semi-continuous and can only treat clarified liquors. Substantial savings could be made if the operation were continuous, thus reducing the resin inventory, and if the costly filtration and clarification steps could be eliminated. Work has been in progress over the past twenty-five years on the recovery of metals from unfiltered ore leach slurries, heap leach liquors and mine drainage waters containing a small amount of suspended solids. In the case of dilute waste solutions, with suspended matter and high flow rates, filtration is usually prohibitively expensive.Continuous ion exchange operation using fluidized beds has made it possible to reduce considerably the resin inventory and to allow the treatment of unclarified liquors and slurries. The subject has been reviewed on several occasions (1. 2. 3, 4). Although fluidized beds in multi-stage columns are an improvement over previous resin-in-pulp treatment, they have complicated internal structure and I or involved automatic control systems to move the resin stages.Slater and Prud'homme performed studies on multi-stage fluid beds, and concluded that the performance of single, deep fluidized beds of resin with continuous or semi-continuous flow of both phases would be worth investigating, particularly in the case of very dilute metal solutionsoi, A pilot plant was designed and built at CANMET in a cooperative project with Dr. Slateroi. The use of deep fluidized beds of resin (with a large height/diameter ratio) for treating dilute solutions has the attraction of engineering simplicity. Extreme simplicity of equipment design is considered of paramount importance when treating, for example, mine waters at high flow rates, possibly of the order of several million gallons per day. This paper describes the development of the concept of single-stage fluidized bed continuous ion exchange arid reviews the pilotplant development of processes for the treatment of base metal mine water, cyanide waste water, uranium leach liquor and uranium mine water."

Jan 1, 1980

By M. K. Walter

"The underground operations of Kidd Creek Mines uses an average of 450 U.S. gpm water during regular operations, and discharges an average of 800 U.S. gpm through consecutive staged pumping lifts of 1200f t.Water is used as a flushing medium for the drilling operations and as a dust suppressant for drawpoint mucking. The water retains an average of 5% suspended solids by weight which is removed underground before being pumped to the surface. This paper outlines the treatment of the mine water for the removal of suspended solids which are chemically treated to settle out as a sludge product of 60% solids by weight. This process reduces the sludge handling operation by 50%. The sludge is systematically disposed into the ore passes.IntroductionThe Kidd Creek minesite, located 27 km north of Timmins, Ontario, produces 4.5 million tonne s of copper, lead, zinc and silver ores annually from its No. I and No.2 underground mining operations. Both mines use a sublevel blasthole stoping method of extraction.Water from the underground operation passes through a series of surface settling ponds, where it is treated to remove heavy metal ions from solution. A blend of the pond effluent with fresh water from Kidd Creek, supply the underground with water at an average rate of 450 U.S. gpm. Water flows from the production zones through coarse sumps and into a main pumping station. There are five main pumping stations underground, three in No. I Mine and two in the No.2 Mine.The No. 2 Mine pumping stations are located on the 4800 level and the 3600 level. All water below the 3600 level to the shaft bottom is controlled from the 4800 pumping station. The water from the 4800 pumping section is combined with water from the remaining levels of the No.2 Mine at the 3600 pumping station. Water from the 3600 pumping station is pumped to the 2400 pumping station in No. I Mine.The No. I Mine pumping stations are located on 2800, 2400, and 1200 levels. All water below the 2400 level to the shaft bottom is controlled through the 2800 pumping station. This water is pumped to the 2400 pumping station. The 2400 pumping station also controls the water below 1600level to 2400 level."

Jan 1, 1988

By D. François-Bongareon

"In previous publications by the author, significant advances and generalizations of the theory of sampling of broken ores, prompted and confirmed by experimental results, have been presented. This paper shows how to implement these theoretical results in practice. The models to be used are reviewed, with a description of the various sampling parameters involved. Then the methodology for the experimental assessment of those parameters, including segregation factors, is described in simple and practical terms, along with the use of handy sampling protocol charts. Finally, the particularly important case of processing data from a sample-tree experiment is detailed, and examples of applications covering all situations and types of experiments are provided.IntroductionThe reliability of a sample of crushed material , i.e. the expected (average) deviation between its true grade and the true grade of the lot it comes from, mostly depends upon six factors:• the size (mass) of the sample (and to a minor extent, for small lots, the size of the lot);• the size of the rock particles (degree of comminution)• the nature of the mineralization at microscopic scale (i.e., within even the smallest particles);• the geological nature of the mineralization and mineral of interest;• the grade of the lot in the mineral of interest; and• in some cases, the state of segregation of the lot.In practice, one should first disregard segregation problems, and assess sample variability in the segregation-free case. Once the numerical parameters governing this variability have been assessed, then segregation factors can be assessed in turn. In the case of laboratory standard sample preparation procedures, segregation can usually be disregarded altogether."

Jan 1, 1993

By M. A. T. Cocquerel, J. G. Eacott

"Simulation techniques may be used to optimize existing processes and plant operations, and in plant design at the engineering study stage of a project . This paper describes the latter application, using a computer, to optimize the concentrate stockpile requirements of a modern copper smelter.A large open-pit mining operation is envisaged and it is assumed that copper concentrate is smelted in two flash furnaces. The proposed mining plan gives rise to metaffurgical considerations in smelting which dictate that concentrate blending is carried out in the early years of operation in order to eliminate wide variations in the grade of concentrate charged to the furnaces . Stockpiling facilities are therefore necessary for blending purposes and also to provide a strategic reserve of concentrate to ensure continuity of concentrator and smelting operations.The objective of the study is to determine the optimum size of concentrate stockpile required to take account of both metaffurgical and strategic considerations. Two separate simulation models are developed to achieve this objective. The first model is used to determine the minimum size of stockpile necessary for blending purposes to smooth out variations in concentrate grade between 15 - 60% Cu, with associated 34 6% Fe, respectively, and ensure that the concentrate fed to the furnaces does not faff below 12% Fe, which is considered a minimum for good smeltability.The second simulation model covers a projected three-year operating period and indicates the maximum size of stockpile which is likely to develop during scheduled and unscheduled shutdown periods in the concentrator and smelter.Covered storage volume and conveying facilities for the optimum size of concentrate stockpile can be designed based on the results generated by the simulation models."

Jan 1, 1980