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By P. Fellner, A. Sýkorová, J. Thonstad, J. Jurišová, V. Khandl

Adsorption of S02 on alumina used in the aluminium industry, so-called smelter grade alumina, was studied in the temperature range 15 C-120 C. The specific surface area of the smelter grade alumina was 88 m2/g. It was found that at temperatures lower than 40 C, sulphur dioxide is bonded to alumina reversibly by physical forces, and the adsorption can be described satisfactorily by the Langmuir adsorption isotherm. The heat of adsorption was estimated to be -33 kJ.mor1. At temperatures in the range 80 C- 120 C, which prevail in dry scrubbers in the aluminium industry, the heat of adsorption was determined to be -56 kJ.mor1. When S02 was adsorbed at temperatures higher than 80 C, about 30 % of the S02 could not be desorbed when heating the samples up to 250 C. In the presence of S02 and oxygen, the formation of sulphate was determined at temperatures above 90 C.

Jan 1, 2006

By John F. Vynne

This paper will discuss several of the systems and processes that have been developed over the past few years, and are now being utilized by major mining companies on draglines in North America, Australia and South Africa. These technologies include production monitors and reporting systems, GPS navigation systems, structural integrity monitoring systems, structural analysis studies to increase the dragline RSLs, and inspection and response strategies. The use of these technologies provides significant increases in dragline production, and reduces maintenance costs and downtime. The question that needs to be asked is what value is there in monitoring? Increasing dragline productivity is dependent upon to two items with the overall goal being to maximize the number of bank cubic meters moved and to uncover more coal: 1) maximizing payload and 2) maximizing the number of cycles. If you do not measure and try to optimize the parameters that impact payload and number of cycles, there is very little chance that you can improve overall dragline productivity. These parameters are: Maximizing payload is dependent upon the following: ? Geology ? Blasting ? Monitor weighing accuracy ? Engage location ? Disengage Location ? Rigging ? Suspended load ? Bucket characteristics ? Operator skills ? Minimizing rehandle Maximizing the number of cycles depends upon maximizing digging time, and minimizing cycle and delay/down times by: ? Minimizing Operational Delays ? Minimizing Mechanical Delays ? Minimizing Idle time ? Minimizing Walking time ? Minimizing Fill Time ? Minimizing Swing Angle ? Maximizing Swing & Return Rates ? Minimizing Spot Time A third reason to monitor, conduct structural analysis studies, and have inspection strategies is to prevent catastrophic failures such as shown in the following photographs (figures 1 & 2).

May 1, 2008

By Otto Engel

"A multipurpose program for a hand-held programmable TJ-59 calculator and PC-100C printer facilitates the evaluation of various mineral processing tests and the calculation of mass and metal balances. Its application is demonstrated by some example problems. Detailed user instructions are given for each application.IntroductionThe program listed in Appendix I is written for a programmable Texas Instruments TI-591 PC-100C calculator I printer combination. The program is applicable for the evaluation of:1) screen analyses;2) flotation test s; and3) mass and metal balances.Each application is demonstrated by an example problem and corresponding detailed user instructions. The ""measured"" data in the examples are fictitious and do not represent a real operation.The steps in using the program are the same for each of the three types of application, viz.:1. Enter the number of product fractions and press ""A"".2. Enter the measured weights (0) of each fraction and press ""RIS"" each time. After the last fraction has been entered, the printer automatical1y displays the total weight.3. Press ""B"" to calculate the weight distribution (Y). Once ""B"" is pressed, the printer automatically displays the weight distribution of all fractions.4. If the accumulated weight distribution (EV) is also needed, press ""RIS"". The printer now displays these data. This step is optional and may be bypassed.5. Press ""2nd A"" to prepare for entering of the measured metal assay (M).6. Enter the measured metal assays of each fraction and press ""RIS"" each time. If analytical data exist for more than one metal, each data set has to be entered separately in one sequence.7. Press ""C"" to determine the ""calculated"" head assay (MA) of all fractions.8. Press ""RIS"" to calculate the metal distribution (R). Once ""RI S"" is pressed, the printer automatically displays the metal distribution of all fractions. This operation is optinal and may be bypassed if desired.9. Again press ""RIS"" to obtain the accumulated metal distribution (ER). This operation is optional and may be bypassed.10. Press ""D"" to calculate the gangue distribution in the concentrate (W). One ""D"" is pressed, the printer automatically displays the data for all fractions. This operation is optional and may be bypassed.11. Press ""RIS"" to obtain the accumulated gangue distribution (EW). This operation is optional and ma y be bypassed.12. For more than one metal, press ""E"" to clear the memory for storing the assays of the next. Then press ""2nd A"" and start entering the assays of the next metal as described previously (6 to 11). Steps 5 to 12 may be repeated as many times as desired."

Jan 1, 1982

By P. MacPhail, G. S. Dobby, I. M. Flint

"The objective of the laboratory work described in this paper has been to develop a method f or producing small gas bubbles at moderate to high gas rates in a flotation column. Small bubbles are particularly important for attaining reasonable flotation rates of fine particles. An ultrasonic aerosol generator has been employed to add frother solution directly to the gas stream. Using porous glass and filter cloth spargers, aerosol frother addition has produced bubbles at 1/2 to 2/3 the diameter of those produced by conventional frother addition to the liquid, at a lower frother consumption.IntroductionGas bubbles are generated in a flotation column most commonly by forcing gas through a porous medium called a sparger. Typical sparger material includes rubber (pierced with small holes) , fabric (such as filter cloth), fritted glass, or porous metal (steel or bronze) . Solid sparger material is suitable for laboratory use; however, for industrial applications a flexible sparger such as rubber or fabric is necessary to prevent the eventual plugging of orfices with solid particles and/ or precipitates. An exception to this is the use of a gas liquid contactor external to the column in combination with pressure release of gas in water, a method that has been demonstrated to produce very small gas bubbles. The typical objective of bubble generation for flotation column purposes is to produce bubbles with diameter of about 0.1 cm (and less) at a gas superficial velocity of 1 cm/s to 3 cm/s. In general, the smaller the bubble, the better. In man y cases, though, there is little , or nothing, gained by generating bubbles less than about 0.05 ern in diameter.Several factors affect bubble size when generated by a sparger: gas rate, orifice diameter, sparger surface area, surfactant type, surfactant dosage, and static head. At very low gas rates bubble diameter is independent of flow rate and is proportional to orifice diameter. At the moderate gas rates encountered in flotation columns bubble diameter, db' is related to gas flow by: db oc vg n, where vg is superficial gas velocity (rate of gas flow per unit column area). For porous steel and filter cloth, n has a value of about 0.25 for vg between 0.2 cm/s and 2 cm/s(2). The magnitude of the effect of orifice diameter at these gas rates is unclear. The effects of frother type and concentration on bubble size issuing from a single orifice have been measur ed by Miagkovaru , shown here in Figure I(IS). The reduction in bubble diameter upon surfactant addition is due to the lowering of interfacial tension created by adsorption of surfactant at the gas liquid interface."

Jan 1, 1988

By Karl J. C. Harries

"9.1. INTRODUCTIONIn Chapter 5 of Volume I royalties are dealt with and several topics would appear to be duplicated in this Volume. However, Volume I may be looked upon as a general guide, whereas this Volume is a reference volume. It deals in considerable detail with specific provisions and concepts relating to royalties between private parties. This does not, however, mean that Chapter 5 can be ignored. It should be looked upon as a source of general information concerning royalties between private parties and as a negotiating aid intended to give an overview of various topics. You may well find that in its general approach it will lead you to ask a question which, hopefully, will be dealt with in this Volume .It should be remembered that royalty agreements are agreements with a particular focus, and so the concepts set forth in Chapter 1 apply to them. This chapter addresses a few considerations specific to royalties that will usually form part of an agreement rather than a stand-alone agreement.9.2. FORMAT FOR ROYALTY PROVISIONSIn many of the cases where royalties arise, they are adjunct to, or part of, other arrangements. For example, a royalty may form part or all of the consideration to be given to an owner upon purchase of his property, or it may arise if a joint venturer ceases to contribute to the venture and has his interest converted to a royalty.1 It is with such situations that this section primarily deals; if the royalty is contained in an agreement that is singularly, or primarily, concerned with creating it, then it is just a normal agreement.Commonly, by the time the right to receive the royalty arises, at least some of the other provisions of the agreement have been performed and/or become redundant. For example, in an option agreement the royalty will be a continuing obligation of the optionee, usually arising only after the option has been exercised.2 In a joint venture agreement, a royalty will often arise where the interest of a non-participating venturer has been diluted and converted to a royalty, in which case the recipient will usually become a passive party, entitled only to his royalty, and most of the rights that he enjoyed as an active venturer will cease to be effective."

Jan 1, 2003

By B. Van den Bogaert, K. Binnemans, T. Van Gerven

Europium is one of the most critical rare-earth elements due to the limited supply of this element combined with a high demand for red lamp phosphors. Therefore, the recycling of europium from end-oflife lamp phosphors has become increasingly important. Photochemical reduction of europium(III) to europium(II) and subsequent removal from rare-earth mixtures by precipitation as EuSO4 is a very selective and efficient method to recover europium from binary europium/yttrium solutions. It is shown that full separation of europium and yttrium is feasible from laboratory prepared non-equimolar Eu/Y mixtures up to a ratio of 1/20, which reflect the composition of red lamp phosphors. Optimal conditions were determined at pH 4 with isopropanol as a radical scavenger, which is a less harmful and toxic compound than the commonly used formic acid scavenger. Europium removal efficiencies of more than 95% were achieved, with a purity of 98.5% of the EuSO4 precipitate. The irradiance of the used light source, expressed in mW/cm2, has a large impact on the illumination time needed to remove europium from the solution. It is observed that higher irradiance results in a shorter induction time, i.e. a period during which virtually no europium is removed, and a faster removal rate once the induction time has passed. The photochemical separation technique was applied to recover europium from leached industrial red lamp phosphor (YOX) powder, with a Eu/Y ratio of 1/30. Europium recovery of 50% was reached with a purity of 96%, proving the feasibility of this method for the recycling of europium from end-of-life lamp phosphors.

Jan 1, 2016

By M. Oliazadeh, A. Vazirizadeh

"The inclusion of small amounts of silicon, phosphorus and sulphur in iron ores or concentrates can have a profound effect on the steel making process and thus on the final iron and steel products. In nature however, it is inevitable that iron deposits contain unwanted elements. Each element has its own distinct effect on the final steel product. The effect of high silicon is the formation of gray iron. Phosphorus mainly increases hardness and strength, lowers solidus temperature and increases fluidity and cold shortness. Finally sulphur contamination causes cracks in the iron texture when smelting and forms white iron when casting. Overall, the acceptable level of phosphorus, sulphur and silica in the iron concentrate should be less than 0.04%, 0.15 % and 1.5 % respectively.Scarcity of high grade iron ores necessitates using lower grade ores with higher amounts of impurities. The purpose of iron concentrators is to increase iron grades to an appropriate level suitable for steel making as well as removing impurities. Although many impurities can be traced in the ore, the most predominant impurities are limited to one or two elements. In this paper the beneficiation methods for removing SiO2, P and S are presented, including flotation, and the parameters that need to be considered in the design of beneficiation plants are discussed. Finally industrial cases are discussed along with their metallurgical performance.INTRODUCTIONIn metal sector, value of iron ores and concentrates is determined by iron percentage as well ascontent of:• Useful elements (Mn, Ni, Cr, V, Ti): natural alloying elements in steel• Deleterious impurities (S, P, As, Zn, Pb, K, Na): impair metal properties or complicate operational process in blast furnace• Fluxing components (Si, Ca, Mg,, Al)Sulphur and copper result in red-short, phosphorus is the cause of cold brittleness and arsenic and copper decrease welding characteristics. Sulphur should not exceed 0.15 in commercial ore. Acceptable sulphur content in ores and concentrates for producing sinter and pellets is 0.6%max, as sintering and heat hardening of pellets remove sulphur 60-90%. Phosphorous content in ores, sinter and pellets is limited to 0.07%-0.15%. 0.00.1% As max, 0.1-0.2% Zn max and up to 0.2% Cu are allowed in iron-ore part of the blast furnace burden for the production of standard conversion pig iron. Slag-forming components are divided into basic oxides (Ca, Mg) and acidic (Si, Al). Higher ratios of basic to acidic oxides are preferred since this reduces raw fluxing material."

Jan 1, 2016

By Nick Vagenas

"Loading and transportation are important activities of the mining process. Automation of these activities in underground hard rock mining has been discussed for a long time. This paper deals with the performance evaluation of a prototype automatic Load-Haul-Dump (LHD) vehicle (a 7-ton GHH LF LHD model) operating in Zinkgruvan mine, Sweden. The purpose of the data analysis is to evaluate the operational and maintenance characteristics of the automatic LHD system in the mine. Graphical and numerical methods have been used for the statistical analysis of the data. IntroductionDuring the past years, the idea of integrated automation and tele-operation on Load-Haul-Dump (LHD) machines (Fig. 1) has attracted considerable interest in the Swedish mining industry. This is mainly due to the demand for improvement in the working environment and due to the rapid development of economic and reliable microprocessors for mining vehicles. Several advantages can be foreseen with the application of a Remote-controlled/Automatic LHD (RAL) system in underground mining. Some of them are (1)':• better utilization of working hours;• less interruption in production at shift changes;• better operational control of a fleet of vehicles;• improved safety; and• one operator may be able to operate more than one vehicle simultaneously."

Jan 1, 1993

By Edouard Masengo

A long dyke will be constructed in a lake with rockfill placed by end dumping method. Its crest will be at elevation 521.5 m compared to the nominal lake level at 520 m. In the first portion, its height will vary from 1.5 m on a 35.1 m thick tailings layer to 18.3 m on about 0.5 m of tailings. In the second portion, the dyke will be built on native silty clay and will vary from 18.3 to 15.2 m in height. The tailings consist in stratified layers of fine-grained soils showing lateral and vertical segregation. The tailings are compressible and may develop excess pore water pressure on loading. The tailings overlie a thin layer of relatively soft native silty clay. The till beneath the silty clay is considered incompressible compared to the material above. After construction, as the tailings and silty clay consolidate, pore water pressures will dissipate and dyke stability will improve. Since dyke stability will increase with time, the primary concern is its stability during construction. As a result of construction activities, three potential modes of failure were identified and recommendations were provided regarding the measures that should be undertaken to increase dyke stability during construction.

May 1, 2003

By Ron Davidson

"The Utility Mine in southeast Saskatchewan is owned by Saskatchewan Power Corporation and operated by Mana/ta Coal Ltd. Up to 3.3 million tonnes of lignite is supplied annually to Saskatchewan Power's Boundary Dam Generating Station. The coal occurs in two predominant seams which are covered by up to 37 metres of saturated fine-grained overburden. Overburden is removed by two draglines working in two pits. Coal is hauled to the mine mouth power station. Comprehensive drainage and reclamation programs are integral parts of the operation. IntroductionThe Utility Mine is owned by Saskatchewan Power Corporation and operated by Manalta Coal Ltd. It is located in southeastern Saskatchewan, approximately 200 km southeast of Regina and 8.5 km southwest of Estevan (Fig. 1). The climate of the Estevan area is semi-arid with long, cool winters and short, hot summers. The mean annual precipitation is 42 cm and the mean summer temperature is 20°C; mean winter temperature is -10°C. The Utility Mine supplies coal to the adjacent Boundary Dam Generating Station. Increases in the station's generating capacity during the past decade have resulted in an extensive expansion program for the Utility Mine. This expansion program, the geological conditions present in the area, and the mining/reclamation operations will be highlighted in the following sections<1.2). History The Estevan area has a long history of coal mining. In the early years, coal was mined via entries into coal seam outcrops along the slopes of the Souris River valley. Surface strip mining was introduced in the late 1920s and progressed to the point where it totally replaced underground mining by the 1950s. Surface mining methods are currently used exclusively by all of the Estevan area operators."

Jan 1, 1985

By W. G. Spencer, E. Sulis, P. M. Tiley

"Dome Mine&apos;s No.8 shaft is located about 1 km northeast of Dome&apos;s No.3 shaft and mill complex in South Porcupine, Ontario.No.8 shaft was brought into production in 1984 to hoist ore and waste for the complete mine as well as to service the mine from the 1600 level and below. Access above 1600 level is provided by No. 3 shaft.The rectangular shaft has six compartments consisting of two skips, a cage, counterweight, manway, and a service compartment. The over-all size of the shaft is 14 ft 7 in. by 20 ft by 7 in. (Fig. 1). A 7th compartment for shaft deepening is incorporated in the bottom 500 ft of shaft. The shaft is 5470 ft deep. The loading pocket is located at 5300 ft and the dump 61 ft below the collar.Surface facilities consist of a concrete headframe and a structural steel hoistroom which houses the cage and production hoists . The service hoist is a 12-ft diameter double drum, with one 1750 HP DC motor, and the production hoist is a 15-ft diameter double drum with twin 1750 HP DC motors.Both hoists have a single clutch and a single reduction gear train built by Orenstein & Koppel, and equipped with Canadian General Electric drives. The cage hoist is equipped with a disc pinion brake.The skips were originally aluminum 14-ton capacity, Rolla-chute design with steel liners on three sides and 4 in. of rubber on the back, to absorb the loading impact. The original guide rollers consisted of fixed 10 in. pneumatic tires. The face-to-face units were eventually replaced with larger airplane-type tires mounted on a spring loaded pivot arm."

Jan 1, 1993

By A. Gillis

"In order to obtain the equipment necessary to automate the equipment it was decided that the Department of Mines and Milltronics would jointly survey the circuit. This survey involved a metallurgical study by the Department of Mines and Milltronics took sound recordings during manual operation, checking in circuit bottle necks and determining the number of sensors that could be applied.The conveyer is driven by 2 PIV drives, each has a variable speed of 3/ 1. The cross conveyer that feeds directly into the rod mill feed chute will contain an automatic sampler and will eliminate the necessity of stopping the belt feeders to take samples. At present, hourly samples are taken from this point and the circuit is upset and cycling results.The Adequate scale is very close to the rod mill to totalize and thereby give a continuous indication and the second purpose is to take a signal from it and control the solution addition to the rod mill.A pneumatic positioner was obtained and put on the scale which produces a 3 tp 15 psi signal which is proportional to the actual tonnage going across the scale. This signal was fed to a bias loader which was mounted on the control panel. From there the operator can set the amount of solution that is wanted by the rod mill. No matter what the feed rate is, the signal coming from the bias loader will follow it automatically. This in turn goes to the pneumatic solution valve (1 ½” ) and the solution goes in with the feed to the rod mill. This is only part of the solution required for dilution since the rest of it is used as a force jet for the feed to the rod mill."

Jan 1, 1965

By P. E. Weidmark

"PART I - Published PapersBallivy, G., and Dupuis, M., ""Laboratory testing for the design of grouted rock anchors""; Proc 6th Pan Am Conf Soil Mech Found Eng; Lima; 1:277-291; 1979.Ballivy, G., Ladanyi, B., and Gill, D., ""The effects of moisture content on the strength of fissured rocks""; Proc 1// 1ntnl Congr 1.A.E.G.; Madrid; //:2:7-16; 1978.Baumgartner, P., and Stimpson, B., ""Development of a tiltable base friction frame for kinematic studies of caving at various depths""; Int J Rock Mech Min Sci Geomech Abstr; 16:4:265-267; 1979.Bawden, W.F., and Roegiers, J.-c., ""Discussion and interpretation of the mathematics governing radial flow in a single horizontal rock fracture""; ISBN 07727-7002-6; Dept Civil Engng publication 79-11; Univ Toronto: 134p;1979.Chatterji, P.K., Smith, L.B., and Insley, A.E., ""Construction of Saline Creek Tunnel in Athabasca oil sand""; Can Geotech J; 16:1:90-107: 1979.Coates, D.F., and Gyenge, M., ""A case history of pit slope design""; Proc 4th ISRM Congr; Montreux, Switzerland; Sept 1979; 1:Theme IV:591-596; 1979.Crawford, A.M., and Curran, J.H., ""A dynamic direct shear machine for determining the time-dependent behaviour of rock joints"": Dept Civil Engng publication 79-08; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A displacement discontinuity approach to modelling the creep behavior of rock and its discontinuities""; Dept Civil Engng publication 79-09; Univ Toronto; 1979.Crawford, A.M., and Curran, J.H., ""A higher-order displacement discontinuity element""; Dept Civil Engng publication 79-10; Univ Toronto; 1979.Cruden, D.M., ""A computer-based system for rock mass description""; Proc 1// Intnl Congr I.A.E.G.; Madrid; //:2:155-159; 1978.Cruikshank, D.J., and Curran, J.H., ""A.B.LE. approach to modelling of fluid flow in fractured rock masses""; Proc 7th Cdn Congr Appl Mech; May 1979; Sherbrooke, Que; 887-888; 1979."

Jan 1, 1980

By P. E. Weidmark

"PART 1-Published PapersAllcott, G.A., and Archibald, D.E.; ""Description of pillar behaviour at Heath Steele Mines""; Can Min Metall Bull; 74:834:80-87; 1981.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of rock Mechanics; 3:32-33; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:37-38; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:38; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses"" ; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:41; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:42; 1979.Baar, C.A.; Discussion of Theme 1: ""Physical properties of intact rock and rock masses""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:46; 1979.Baar, C.A.; Discussion of Theme 2: ""Tectonophysics""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:60; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3: 116; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:116-117; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:117; 1979.Baar, C.A. &apos;; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:118-119; 1979.Baar, C.A.; Discussions (2) of Theme 4: ""Underground openings""; Proceedings 3rd Congress lntnl Soc of Rock Mechanics; 3:119; 1979.Baar, C.A.; Discussion of Theme 4: ""Underground openings""; Proceedings 3rd Congress Intnl Soc of Rock Mechanics; 3:120-121; 1979."

Jan 1, 1982

By J. C. O'Rourke

"Gibraltar Mines Ltd. has been operating its open pit copper mine for fifteen years. In that time over 140 million tons of waste and low grade material has been placed on the mine waste dumps. Laboratory and in-situ testing indicated that enough leachable copper was recoverable from the dumps to justify a leaching operation which would include the construction of a solvent extraction-electrowinning plant. The plant, which was commissioned in October, 1984 was designed to process 3,000 usgpm of dilute copper solution to produce 5,000 tons per year of high quality cathode copper.This paper deals with the start up and initial year of operation of Gibraltar's SX-EW plant. The three areas: leaching solvent extraction and electrowinning are presented with emphasis on the innovations which were made to help overcome the problems unique to die Gibraltar operation.THE LEACHING CIRCUITThe mines waste dumps are leached to produce a pregnant copper solution as feed to the SX-EW plant. To achieve this the dumps are sprinkled with a mildly acidic solution which percolates through the dump. In travelling through the dumps the solution contacts the surfaces of the liberated copper minerals and disolves the soluble copper into the form of copper sulphate. The leaching reaction is achieved by a combination of bacterial action to oxidize the suphides and solublization of the copper into the acid solution.For the bacterial action to occur the dumps must be at or near the ideal pH of 1.9 - 2.1 and the necessary elements of the reaction must be present. Because a bacterial reaction is involved there is a lag time between the time when the conditions for bacterial growth are achieved and when the bacteria are in their active growth phase. Laboratory tests with Gibraltar ore indicated that bacterial action takes two to three months to reach its maximum rate. See figure 1.The expected lag time for the planned rate of bacterial leaching prompted Gibraltar to start irrigating the dumps in late June 1986 for a plant start up in early fall. The lag time proved to be longer than expected and the dumps took until November to start producing copper at the expected rate. See figure 2."

Jan 1, 1988

By G. A. Gillett, C. F. Bonney

"The Davy McKee combined mixer-settler (CMS), is a liquid-liquid contactor, which has been commercially applied to the recovery of uranium from dilute sulphuric acid-bearing leach liquors. CMS differs from most other solvent extraction systems in that the mixing and separation of the phases is carried out in the same vessel.IntroductionTypical solvent extraction plants for the recovery of uranium operate at organic to aqueous feed flow ratios of 1 : 2 to 1 : 30. Because it is beneficial to mix the organic and aqueous phases at an organic/aqueous (O / A) ratio of between I : 2 and 2 : I there is incompatibility between the over-all and operating O /A ratios.Conventional mixer-settler designs require an external recycle of the minority flow which results in the possibility of over-mixing, increased settling requirements and entrainment losses. Other workers have approached this problem in terms o f enhanced sett ling and this has been partly successful.Davy McKee&apos;s approach involved a radical rethink of the principles of liquid-liquid contact and separation.Description of CMSThe CMS has the important advance that the required mixing O/A ratio can be achieved without external recycle and is not affected by the O /A ratio of the feeds. Three zones co-exist within the CMS unit-a central mixing zone and upper and lower separation zones. A pump impeller on a draught tube is installed within the mixing zone and baffles are installed above and below the mixing zone. Conventional mixer- settlers have several undesirable features which ca n largely be eliminated using the CMS system:"

Jan 1, 1983

By Laxman M. Amaratunga, Glen G. Hein

"Most mining operations utilize the tailings coarse fraction for backfilling while the fines have been traditionally disposed on surface in tailings ponds. Recently, paste fill has been developed which uses a fraction of the fine tailings (10-30% w/w of <45 µm) in the fill (Brummer and Moss, 1991). This development allows a mining company to use a portion of the fines for backfill, but still requires surface impoundment of the unused tailings. Reactive tailings have the added problem of generating acid after being oxidized in the presence of air and water. Treating acid mine drainage (AMO) from these ponds has proven to be a costly environmental problem and a long term liability.One practical method to reduce the environmental impacts associated with the tailings pond is to modify the physical and chemical prope1ties of the fine tailings so that they can be utilized in an environmentally safe manner. This was accomplished by agglomerating the tailings into pellets and using theses pellets as a coarse aggregate in a total tailings paste fill. This technology does not require any supplemental sand or gravel to be added to the fill.This paper presents the ongoing research into utilizing these tailings agglomerates as a coarse aggregate fraction, in order to create a total tailings paste fill which has supe1ior strength characteristics.IntroductionThe Canadian mining industry produces more than 500 million tonnes of tailings and waste rock annually. The waste rock is usually returned underground for use as rock fill in underground support with the tailings remaining to be disposed of in tailings ponds. Tailings consist of a coarse and fine fraction, and can be either reactive (acid generating) or non-reactive. In an effort to minimize tailings disposal, the coarse fraction has been used underground in backfill along with alluvial sands. This diversion of the coarse fraction, however still requires that the fine fraction be disposed of in tailings ponds. This disposal creates the potential for acid mine drainage (AMO), since in the presence of air and water, reactive sulphide tailings can generate acid. It is therefore an important goal in the mining industry to minimize the disposal of fine tailings into tailings ponds since the treatment of AMO is a costly and long-te1m environmental liability."

Jan 1, 1997

By C. Díaz, A. D. Dalvi, C. E. O'Neill

"Roast-reduction smelting (RRS) of nickel concentrate, combined with flash smelting of copper concentrate, was one of the process options that Inco considered as a means of meeting its 1994 sulphur dioxide abatement target at its Copper Cliff operations. RRS involves deep roasting the nickel concentrate in fluid bed roasters. The roaster off-gas is treated in a sulphuric acid plant, and the low sulphur calcine is reduction smelted with coke in an electric furnace to yield a sulphur-deficient matte. The latter material is converted to Bessemer matte in Peirce-Smith converters, with minimal evolution of sulphur dioxide, and the converter slag is reverted to the electric furnace. Excellent nickel and cobalt recoveries (Ire obtained. This paper discusses the development of RRS, including extensive commercial scale testing at Inco&apos;s Thompson Smelter, and its key operating parameters. IntroductionReduction of sulphur dioxide emissions has been a long-standing endeavour at Inco. At the company&apos;s Ontario Division, as shown in Figure 1, sulphur dioxide emissions were reduced from about 2 million to 685 000 tonnes per year over the period 1970-1985. Most of this impressive improvement was the result of a sustained research effort which led to increased rejection of pyrrhotite from the ore during milling while maintaining an acceptable nickel recovery. Between 1970 and 1985, the combined grade of the separate nickel and copper concentrates feeding the Copper Cliff Smelter increased from about 10070 to about &apos;lIJ0J0 Cu +Ni and the combined S/(CU +Ni) ratio decreased from about 2.8 to 1.65."

Jan 1, 1994

By Robin N. Adair

The Halfmile Lake stratiform volcanogenic massive-sulfide deposit consists of pyrrhotiterich breccia-matrix sulfides and pyrite-pyrrhotite-rich layered sulfides that form a laterally continuous sheet stratigraphically underlain by an extensive pyrrhotite-chalcopyrite-quartz stringer zone. Mineralization is hosted by a sequence of felsic pyroclastics and volcanic-derived, fine-grained metasediments. The proportion of volcanic to sedimentary rocks indicates increasing vent proximity from southwest to northeast across the deposit area. Fine- to medium-grained felsic pyroclastics and roughly equal amounts of fine-grained metasediments comprise the stratigraphic footwall to the mineralization. Quartz-feldspar porphyritic bodies cut this sequence and may have provided the heat source for exhalite activity. Felsic tuffs, locally crystal-rich, and lesser amounts of metasedimentary rocks make up the immediate stratigraphic hanging wall. This sequence is stratigraphically overlain and locally interfmgered with mafic volcanics that contain substantial proportions of felsic fragments. Deposition of sulfides post-dated the beginning of felsic volcanism, which places the deposit in the middle of the Lower Tetagouche Group. Compositions of the volcanic rocks range from basalt to rhyolite and exhibit sub-alkaline affinities. Strongly overprinted whole-rock, trace-element, and REE geochemical signatures resulting from mineralizing hydrothermal activity confirm the genetic relationship between the stringer zone and the mineralized sheet. Large-scale D1 folding has overturned the entire sequence. D1 thrust faults and related sheath-like folds have profoundly affected sulfide distribution. D2 is manifested by a penetrative S2 foliation, and D3 is represented by large-scale open folds.

Jan 1, 1992

By D. W. Heard

Electrospark Deposition (ESD) is a low heat-input, high energy density, micro-welding process. The short pulse duration and high-pulse frequency, in combination with the small amount of material transferred during each pulse, results in the evolution of extremely rapid cooling rates, believed to approach 105-106 °C/s [1]. This rapid solidification event results in the refinement of the microstructure to the nanostructured or amorphous level [2, 3]. Figure 1 displays the resultant microstructure of an Al-17%Si alloy deposited by the ESD process. Furthermore, it was determined that the microstructure of the deposit remained unchanged after subsequent layers were welded onto the presented layer. Freeforming of materials is a rapid prototyping process in which the deposited feedstock is built-up to the desired geometry in a near-net shape manner. Figure 2 displays the automated set-up employed produce three dimensional freeformed objects via multiple deposition trials of the ESD process. The automated XYZ table is equipped with CNC software to transform CAD renderings into three dimensional tool paths. Figure 3 displays an Al-12Si freeformed gear (2" diameter, 12 teeth) produced by the ESD process, prior to removal from the substrate. An SEM analysis of the freeformed component showed that the microstructure did not coarsen during the deposition of subsequent layers during the freeforming process, as depicted in Figure 4, resulting in the production of a bulk nanostructured component. This paper represents a summary of the capabilities of ESD to fabricate advanced materials possessing refined microstructure. In particular, results on the microstructural evolution obtainable for various wear-related systems will be presented. In addition, the feasibility of using the ESD process to freeform three-dimensional objects via multiple deposition passes will be presented and accompanied by a characterization of the properties of the freeformed component. [1] R. N. Johnson, "ElectroSpark Deposition: Principles and Applications," in Society of Vacuum Coaters, 2002, pp. 87-92. [2] J. Milligan, D. W. Heard, and M. Brochu, "Formation of nanostructured weldments in the Al-Si system using electrospark welding," Applied Surface Science, vol. 256, pp. 4009-4016. [3] S. Cadney and M. Brochu, "Formation of amorphous Zr41.2Ti13.8NilOCul2.5Be22.5 coatings via the ElectroSpark Deposition process," Intermetallics, vol. 16, pp. 518-523, 2008.

Jan 1, 2010