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By M. R. KEYS

"The property is about 120 miles north of Red Lake in the Favourable Lake area, district of Kenora ( Patricia part) , Ontario. The original discovery was staked by K. Murray in December, 1927. The first development work was done by the Favourable Lake Mining and Exploration Company. In 1936 Newmont Mining Corporation acquired control of the property and developed it to production by September, 1939. Production to the end of 1945 had a value of $7,534,656 from 441,457 tons of ore, an average value of $17.04 a ton . This production was mainly from gold and silver, with lead and zinc of subordinate importance.The rocks in the vicinity of the mine trend in a north-south direction. From east to west the sequence is: (1) A dark-coloured, massive rock, presumably a flow, composed mainly of quartz and biotite. ( 2) A buff- coloured, scricitized rock locally containing quartz phenocrysts and black hornblende. This is an alteration product of the dark variety and is later than the folding. It occurs mainly within the dark rock. (3) A series of sediments comprising thin beds of greywacke, grit, quartzite, conglomerate, and iron formation. ( 4) A gabbro sill, now largely altered to amphibolite.With the exception of the buff type, these rocks dip 80°E. The buff rock has a very irregular outline in comparison with the other rocks and appears to dip somewhat more flatly east. Felsite and diorite dykes are also present. These have a general strike of S.60°E. and dip 65°S .W. A post-ore diorite dyke trending N.15°E. occurs about 2,000 feet east of the mine area."

Jan 1, 1954

By M. J. Merwin

Heat treated seamless tubular products are processed in a fashion that can be easily modified to produce dual phase microstructures, through the implementation of intercritical soaking prior to quenching. Dual phase steels are known to possess low yield strength compared to tensile strength. The steel line pipe market is interested in low yield-tensile ratio materials for application in reel barge pipe laying operations. In the current work compositions of commercially produced material already in use for the line pipe market were examined and dual phase microstructures were successfully developed. The microstructure and properties of the materials were found to be stable over the range of intercritical temperatures employed. Tempering temperature and, somewhat unexpectedly, small chemistry differences were found to have the largest influence on properties. Excellent toughness performance was observed in samples tempered at temperatures in excess of 550°C. Tempering in this temperature regime produced materials with strength suitable for application as API X60 line pipe. A difference in tensile strength between steels of similar chemistry was observed, and is not fully understood at this time. Welding performance was considered through simulation of three regions of the heat-affected zone. Neither intercritical temperature, nor tempering temperature, were found to have a significant influence on the heat-affected zone properties. The peak temperature achieved in the thermal cycle was the only significant factor. The hardness of the heat-affected zone was found to increase continuously with increasing peak temperature. While toughness was found to be low for both partially-transformed and coarse-grain heat-affected zones, the toughness of the fine-grained heat-affected zone was substantially better. Based on the results found in this laboratory study, a limited production trial is expected.

Jan 1, 2005

By J. M. Cummings

The importance of controlled sand properties to the production of quality castings has been recognized by foundry men for many years, but the maintenance of requisite standards results in a relatively high sand consumption, which in steel foundries, for instance, may range from half a ton to over one ton of sand per ton of finished casting. During the war, obtaining and maintaining adequate sand supplies became a serious problem in many foundries, owing to increased wartime production and serious 'bottlenecks' in transportation. As a result, considerable interest was aroused in the possibilities of various sand reclamation systems by which up to 85 or 90 percent of discarded sand may be returned to a usable condition. During the past five years or so, a number of sand recovery installations have been made by both ferrous and non-ferrous foundries in Canada and the United States, and have resulted in substantial reductions in sand costs in every case. The possible benefits of sand reclamation are particularly apparent in British Columbia where no suitable sands for steel and heavy iron foundries occur naturally, and where freight costs on imported sands are considerably higher than initial selling prices. This is well illustrated by the fact that silica sand, which sells for about $2.00 per ton at pits in Illinois, is laid clown in British Columbia for around $11.00 or $12.00 per ton. The operating costs quoted for certain of the reclamation schemes, i.e. $0.25 to $1.50 per ton, may actually be no higher than the cost of waste sand disposal in city-located foundries. In view of the foregoing considerations, the British Columbia Research Council has undertaken a study of the possibilities for sand reclamation, with particular attention directed to its application in local steel foundries. No commercial occurrences of silica sand are known in the Province, and, with this in view, the Council previously investigated the possibility of recovering silica sand from local impure sands. Results were definitely encouraging, but it was felt that the present market is too small to permit economical production by the method developed.

Jan 1, 1947

By Walter Greaves

THIS paper describes a series of tests performed at the Hollinger mill during the past four years, in which both metal and rubber-lined pumps were used for pumping sand and slime. While conditions will not be the same at other mills, it is felt that the information gained from the tests should be of value to other operators and of assistance to them in selecting the most efficient size of pump and kind of lining for economical operation. For the successful operation of pumps with rubber linings there are certain conditions, which, if recognized at the outset, will save the operator a good deal of work and trouble. First is the necessity of keeping sharp pieces of steel, or anything that might cut the lining, out of the pumps. This can be effected by placing screens in the proper locations. Second is to recognize the fact that the rubber lining of the shell, and the rubber cover of the impeller, are subjected to very different conditions. The shell lining, :being at rest and well supported by the metal case, has a relatively easy job compared ? with the impeller covering, which is subjected to centrifugal force and has a tendency to tear itself away from the metal core. Also, the pulp travels through and past the impeller at a high velocity. These two factors, centrifugal force and velocity of the pulp, would appear to set a limit to the speed at which, for a given set of conditions, it is economical to run a rubber-covered impeller. Therefore, a pump should be large enough for the work to be done. To get the best mechanical efficiency, it must be operated somewhere near the limit of its capacity, but this may not give the best overall cost, because the saving in power, due to the high efficiency, may easily be lost through excessive wear on the lining. It would appear that this condition is more critical in a rubber-lined pump than in one lined with metal.

Jan 1, 1939

By G. E. Joe

"This is a panel discussion on the milling experience with large volume flotation machines in Canada. Over the past ten years, the Canadian milling industry has been characterized by treatment of ever increasing tonnages of lower grade ores. In order to treat these ores efficiently, mineral processors have been employing increasingly larger volume flotation machines. In older established mills, standard flotation cells have been replaced by the larger volume models with very apparent savings in mill floor space as noted in the Opemiska copper mill.We are pleased to hear from four mill operators with extensive experience on large volume flotation machines. They are:1. I. E. Wahn, metallurgist with Falconbridge-Manibridge, who will describe their mill experiences with the Tank cell and the Maxwell cell.2. G. M. Hughes, mill superintendent of Sherritt Gordon-Ruttan, will detail recent experience with the Agitair cell.3. J. R. Maxwell, mill superintendent of Opemiska Copper, will report on their Maxwell cell installation and operation.4. A. H. Winckers, superintendent of Development and Technical Control, Sullivan Mill, will present a paper written by H. J. Chalmers describing Cominco's past experiences with large volume flotation machine installations including Denvers, Wemcos and Maxwells."

Jan 1, 1974

By W. Qin

The nucleation of d hydride in Zircaloy-4 has been studied using electron backscatter diffraction. Hydrides prefer to nucleate at intergranular rather than intragranular sites. (0001) a -Zr//(111) d -ZrHx is the most common crystallographic relationship. The hydride platelets are generally interlinked each other and form a chain-like configuration. Based on thermodynamic model, the formation mechanisms of the hydrides at different sites and the interlinked structure were analyzed. We found that the misorientation of two adjacent grains has a crucial effect on hydride nucleation sites, and the tensile stress near the tips of hydride platelets is a key factor affecting the appearance of interlinked structure.

Aug 1, 2010

By M. Mohan Rao

A Sievert ' technique was used. The few determinations made on 10 per cent and 20 per cent V alloys did not show any deviations from Sieverts' law up to I>N2 = 1 atmosphere. The same held true for alloys of 0.2 per cent Ti. Nitrogen absorptions in Fe-Ti alloys, containing from 0.50 to 0.77 per cent Ti at 1,600°C, follow Sieverts' law at low PN2 values, but, due to the formation of a delta titanium nitride phase, deviate from this law beyond critical values of PN2? There is evidence that this phase is Ti1.1N. The behavior observed can be described by the followmg equations: Ti,.,N(o) = 1.7Ti + N K = (fT; ? % Ti)1?7 (f;;.? %~_) = 0.0020 K = fN. %~ = 0.042 '\)-:>N 2 I NTRODUCTJON Nitrogen's solubility in liquid iron-vanadium alloys was investi-gated briefly by Brick and Creevy ( 1), and more extensively by Kash-yap and Parlee (2) and by Pahlke and Elliott (7). Nitrogen's solubility in liquid iron-titanium alloys

Jan 1, 1963

By R. O. Loutfy, J. C. Withers

"Titanium powder is produced from a small fraction of sieved sponge, HDH of sponge or gas blown from a melt produced billet. Ore from foreign sources is chlorinated in a very high temperature fluid bed to make TiCl4, then vacuum distilled purified and then reduced to sponge by Kroll processing. Domestic ores of ilmenite or perovskite can be carbothermically reduced to Ti2OC which can be chlorinated at low temperatures of 250–400°C to produce TiCl4 which can be electrolyzed in a fused salt to Ti powder, or the Ti2OC used as an anode in electrolysis to produce Ti powder. The electrolysis process can be operated continuously to produce Ti powder at a lower cost than Kroll processed sponge including produce control size powder. Electrolysis processing was scaled up and operated at the commercial demonstration stage on a continued basis that verifies the potential of a low cost process to produce Ti powder.INTRODUCTIONTitanium and alloys of titanium powder have been elevated to strategic status resulting from a focus of producing meltless titanium components as well as a feed for additive manufacturing processing. However, the high cost of titanium powder has limited wide scale implementation of using titanium powder as a feed to any process to produce titanium components. Titanium powder free of alloying is available as sponge fines from the Kroll process as illustrated in Figure 1. Titanium alloy powder is much more expensive, also illustrated in Figure 1 produced from an alloy ingot or mil product as the feed and then transformed into powder by one of several melt form processes which accounts for the high cost of titanium alloy powder at approximately $60–150/lb. Ton lots of titanium alloy powder (i.e. Ti-6Al-4V) from one source is quoted at $120/lb. in the U.S. It appears that alloy powder production cost by current processing is not sensitive to volume."

Jan 1, 2016

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 Seungwoo Park, Whuisu Shim, Roe-Hoan Yoon, Kaiwu Huang

"A comprehensive froth model has been derived from first principles. It is based on predicting the kinetics of film thinning in the presence of particles of varying sizes and surface hydrophobicity. According to the Reynolds lubrication theory, the thinning of a horizontal film is driven by the excess film pressure (p), which is determined by the capillary pressure (pc) and the disjoining pressure (II). In the presence of particles, macroscopic capillary pressure (pc) is affected by the local capillary pressure (pc,local) in the vicinity of a particle and the number of particles in the film (N1,film), which in turn vary with the hydrophobicity and size of the particles. The Reynolds equation corrected for pc,local and N1,film is then used to predict the time (or froth height) at which a film is thinned to a critical rupture thickness (hcr), which in turn can be predicted using the Vrij’s capillary wave model (1966) after a modification to incorporate the contributions from hydrophobic disjoining pressure (IIhp). The model can predict bubble coarsening represented by the bubble size ratio between the top (d2,t) and bottom (d2,b) of a froth as functions of particle size and hydrophobicity, ? -potentials, frother dosages, froth height, gas flow rate, salt concentration, etc. The model predictions have been verified in experiments.INTRODUCTIONFroth flotation is regarded the best available technology today for upgrading minerals and coal fines. In the pulp phase of a flotation cell, air bubbles collect hydrophobic particles dispersed in it and rise to enter the froth phase on the top. Once in a froth phase, the bubbles coalesce with each other and become larger, resulting in a decreased bubble surface area. Under conditions of limited surface area, less hydrophobic particles drop off the bubble surface and return to the pulp phase. Thus, bubble coarsening provides a useful mechanism of increasing product grades. On the other hand, uncontrolled coarsening results in a decrease in bubble carrying capacity and hence throughput."

Jan 1, 2016

By Anthony McDonald, Dominic Verhelst

"Hudson Bay Mining and Smelting Co. Ltd. (HBM&S) started up a zinc plant and smelter in late 1930, employing technology of that period. The zinc plant has since converted from a roast/leach operation to a two stage pressure leach operation to eliminate sulphur dioxide emissioms. The smelter, however, is still continuing with much of the original roaster/reverberatory furnace technology. This paper reviews the process flow of the copper smelter, with some comments on past and recent changes.Smelter FlowsheetPrior to July 1993, the HBM&S smelter operation consisted of seven areas, namely feed preparation, roasters, reverberatory furnace, converters, anode casting, fuming plant, and dust collection. A simplified smelter flowsheet is shown in Figure 1. The data for tonnage of materials moving through the circuit are presented in Tables I to 4, and are representative of the time period from July 1992 to June 1993. A change in the zinc plant technology in 1993 resulted in the elimination of a zinc-bearing sulphide leach residue from the smelter feed blend, thereby greatly reducing the zinc content of slag produced in the smelter. Consequently the fuming plant and ancillary equipment, such as the drying plant, were shut down. These areas will, however, be described. With the closure of the zinc fuming furnaces all the slag from the reverberatory furnace is now discarded to stockpile.The over-all smelter metal recoveries are 96% for Cu and rn.5010 for Au and Ag. Figures 2 and 4 indicate historical production numbers of copper, silver and gold, respectively. A review of the minor element distribution in the smelter can be found elsewhere."

Jan 1, 1994

By R. M. de Carvalho

Empirical and phenomenological models of SAG and AG mills have dominated the scene for the last few decades, being used nearly every time a mill is scale-up or an attempt is made to improve its performance. However, in spite of their success, these models fail to describe the details of both the grinding environment and the breakage mechanisms. The result is a generally good track record, with a few ? but very important ? instances when these models fall short in their predictions, which can be very costly. Further, nearly all of these models require detailed calibration to industrial and pilot plant data, being costly in their application, although still with limited predictive capabilities. Recognizing these challenges, a new model framework has been proposed which overcomes the limitations of currently used models by coupling the microscale population balance model, the discrete element model and functions that describe internal classification at the mill grate. This model framework has already been applied to describe breakage in ball mills, with showing very promising results. However, in the case of SAG and AG mills a much greater level of complexity is involved in the model, given the interaction between the mechanical environment that is formed totally or, in part, by the ore itself, and the breakage response in the mill. The paper describes the model framework, demonstrating how it can address appropriately effects such as mill speed, mill filling and size distribution, on the response from the mill. Some of the advantages of the model are related to its intrinsic multi-component and dynamic nature, which allows simulating appropriately mixtures of ores of different competences in a transient manner.

Jan 1, 2012

By T. Lu

Microwaves have considerable potential as an energy source in both mineral processing and metals extraction. In this research, the utilization of microwave radiation for the heating of a gibbsite type bauxite ore was investigated. Firstly, thermogravimetric studies were performed in order to characterize the dewatering processes of the ore. Secondly, both the real and the imaginary permittivities, which are the fundamental properties that determine the interaction of microwaves with any given material, were measured for the bauxite ore as a function of temperature. The changes in the permittivities were shown to be related to the dewatering processes. Subsequently, the effects of operating variables such as processing time, sample mass, particle size, microwave power, sample bulk density and carbon and magnetite additions on the sample temperature in a microwave field were studied. The results showed that under suitable conditions the bauxite ore could be readily heated. Therefore, there is considerable potential for the application of microwaves for the heating of bauxite ores.

Jan 1, 2007

By K. Sun

The exothermic behavior of the intermetallic reactions inside an Al-Mn metallic powder compact and its assimilation fundamentals in molten aluminum were examined. In this study cylindrical compacts were manufactured and tested. The experimental work involved simultaneous temperature measurements of three thermocouples as well as detection of the apparent weight of the cylindrical specimen during its assimilation in liquid aluminum. The details of the intermetallic reactions prior to the compact complete disintegration were investigated by microscopically studying the morphology and local composition in a partially reacted compact. The variation of manganese concentration in the melt was determined by chemically analyzing the samples taken from the molten aluminum at specific time intervals. The exothermic reaction increases the compact's temperature above that of the liquid aluminum. In tandem with this exothermic reaction, the compact is suffering a swelling action. This swelling action disintegrates the cylindrical specimen, which is assimilated into the liquid aluminum. The intermetallic reactions proceed in a stepwise manner. Both the bath temperature and the compact composition impact on the exothermicity of the intermetallic reactions. However, when the amount of Mn in the compact is greater than 75-wt%, the maximum centerline temperature rise is slightly reduced as Mn amount increases.

Jan 1, 2004

By Heinrich Möller

The electro-oxidation of 0.1 M ethylene glycol at gold, platinum and gold-platinum alloy electrodes in different heat treatment conditions has been studied in 0.5 M NaOH, at room temperature. Poisoning of all the electrodes occurs. However, the alloy electrodes poison at a faster rate than the pure gold and platinum electrodes. This indicates that the initially higher activity of these electrodes cannot be the result of decreased poisoning through a "third-body" effect.

Oct 1, 2003

By J. F. Dorion

This paper presents the results of on-going investigations on the performance of support systems in Québec underground mines. The focus of these investigations is on site observations complemented by laboratory investigations. Of particular interest is deducing the corrosion rate, at selected mine sites, as determined by the installation of monitoring coupons.

May 1, 2009

By H. R. Kempthorne

"The Property of Bevcon Mines Limited, including a recently acquired adjoining south group, occupies an area of 2,625 acres in northeast Louvicourt township, Abitibi county, Quebec. It is situated 17 miles east of Val d'Or at the junction of provincial highways 58 and 59.Gold was first discovered in the area in 1931 by the late S. B. Jowsey on the present property of Buffaddison Gold Mines adjoining to the west. However, it was not until the winter of 1944-45 that diamond drilling encountered encouraging results on the Bevcon claims. During the period 1946-1948 a shaft was sunk to 1,000 feet and considerable development undertaken from the 500-foot to the 1,000-foot levels. Between 1947 and 1950 approximately 24,000 tons of dump ore and 48,000 tons of ore by stoping were milled at Perron Gold Mines. The favourable results of these tests prompted the decision to construct a milling plant at Bevcon. The mill began operating in July 1952, and presently has a capacity of 800 tons per day. In 1953 the shaft was deepened to 1,600 feet and is currently (1955) being driven to 2,225 feet.Total production of gold to the end of 1954 has been 101,024 ounces from the milling of 697,427 tons of ore of an average grade of 0·163 ounces per ton.GeologyThe Bevcon orebodies occur in a granodiorite stock which has an east-west length of approximately 4 miles and a maximum width of 1 1/2 miles. The property covers the central portion of the stock and its contacts with the Keewatin volcanics. Approximately one-third of the northern contact and one-half of the southern contact are contained within the Bevcon claims. The granodiorite has been cut by acid (feldspar porphyry) and basic dykes trending east-west. A younger diabase dyke, 300 feet wide, strikes northeasterly across the property 1,000 feet east of the shaft.The main ore zone, striking east-west, lies between the more northerly porphyry dyke and the north granodiorite-greenstone contact. It has been extensively sheared, fractured, altered, and mineralized by gold-bearing solutions. It is approximately 2,000 feet Jong from the west boundary towards the diabase dyke. The shaft is centrally located and passes through the south contact of the ore zone at the 500-foot level. The zone increases in width and trends northerly with depth, due to the porphyry and greenstone contacts having average dips of 80 degrees and 70 degrees respectively to the north. Thus on the uppermost or 500-foot level the width is 150 to 200 feet, and at the present bottom level at 1,600 feet the width is 400 to 500 feet.Another gold-bearing zone has been indicated 200 to 400 feet south of the main ore zone by limited surface drilling."

Jan 1, 1949

By G. M. Hughes

"400 cubic foot Agitair Flotation Cells are in use at Sherritt Gordon's Fox and Ruttan Mine Cu-Zn concentrators. The 4 impeller model No. 120, used at the Fox Mine Operation, appears to offer superior agitation and aeration qualities compared to the single impeller model 120A, in use at Ruttan. The advantages of the model 120 are most significant with ores of high specific gravity and coarse grind.IntroductionSherritt is using 400 cubic foot Agitair flotation machines at both the Fox and Ruttan Mine Concentrators.At the 3000 tpd, Cu-Zn Fox concentrator, 18 model 120 machines have been in operation since May, 1969.The 10000 tpd, Cu-Zn operation at Ruttan utilizes 49 model 120A machines. The plant started operation in May, 1973.The 4-impeller model 120 flotation cell has been marketed by Galigher since the early 1960's but the single impeller model 120A is relatively new.Model 120 Agitair -Cell DataThe Model 120 (Dwg. 1) has floor dimensions of 101 x 10' and a lip height of 2~ to 4 feet to provide up to 400 cubic feet of volume. This machine is shown with a froth crowder and stabilizer plates which are optional.Each motor drives two agitators. 20 HP is recommended by Galigher but 25 HP is more realistic with massive sulphides. The SG of ?ox ore is about 3. 6.The Model 120 can handle up to 600 cubic feet of air per cell at 2 psig which requires 5-6 HP at the blowers. Air volume to each impeller is controlled by a valve from a main header.During the first 3 years of operation at Fox there were few mechanical problems with the 120's."

Jan 1, 1974

By A. Mular, T. Bond, G. D. Farnell, J. F. Kingston

"ABSTRACT:-A 60—cm semiautogenous grinding-(SAG) mill was designed -for continuous wet operation in closed circuit with a vibrating screen. The circuit is fully instrumented and automated. Part of the control system incorporates Foxboro SPEC 200 controllers around the pump box to maintain level. The rest of the control system involves an ACTION model BC2 basic direct digital controller that handles all other inputs/outputs and contact closure sensing:·The success of the control system is attributed to the uniqueness of specially designed mass flow metering systems. Various control strategies are possible. Employed to date has been control of a mill load set-point by cascading the deviation from load set-point as the set-point to the fresh feed rate ""controller"". Mill feed water is ratioed to total flow rate of solids fed to the mill. The performance of the circuit is being correlated with the performance of full scale SAG circuits in British Columbia to develop models, to assess the feasibility of scale up and to investigate the effects of design variables on responses such as circuit throughput. 1 INTRODUCTIONWhere ore types are appropriate, semiautogenous grinding and/or autogenous grinding (AG) has replaced conventional grinding. Secondary/Tertiary crushing plants are thereby eliminated so that capital expenditures are less and overall comminution costs can be lower (Barratt 1982). To establish the suitability of an ore for SAG circuits, pilot scale tests with 1.7 to 1.8 M diameter mills are often conducted. At least one test per ore type in a deposit has been necessary. Often, each test involves a large amount (up to 50 tonne) of sample, so that the testwork can be expensive and may discourage the use of SAG or AG circuits. However, MacPherson (1977) has employed a continuous 45 cm dry Aerofall mill circuit closed with screens and an air classifier to determine whether ores are suitable for SAG milling. The amount of sample per ore type is of the order of 60 kg or less. Thus an established SAG mill scale up procedure that employs small circuits is highly desirable.The purpose of this paper is to describe a small 60 cm wet continuous SAG/ AG mill pilot circuit assembled at the University of British Columbia to study the feasibility of scale up, to assess the influence of design variables, to develop mathematical models and to evaluate the effects of alternative control strategies on circuit performance"

Jan 1, 1988

By Hajime Miki, Gde Pandhe Wisnu Suyantara, Shigeto Kuroiwa, Eri Takida, Tsuyoshi Hirajima, Masashi Yamane, Yuji Imaizumi, Mitsuru Sawada, Keiko Sasaki

Selective flotation of molybdenite and chalcopyrite using various oxidation treatments (i.e., ozone treatment, hydrogen peroxide, and Fenton-like reagent) was studied in this work. It was found that ozone treatment produced a depressing effect on both minerals, resulting in a lower separation of both minerals. On the other hand, oxidation treatment using hydrogen peroxide (H2O2) and Fenton-like reagent could deliver selective oxidation on chalcopyrite surface and depress its floatability. However, bench-scale copper- molybdenum (Cu-Mo) flotation tests show that oxidation treatment using H2O2 required a higher dosage of H2O2 and a longer treatment time in order to deliver comparable flotation results to that of conventional Cu- Mo flotation using NaHS treatment. On the contrary, Fenton-like reagent could deliver selective Cu-Mo flotation results comparable to that of NaHS treatment at a lower dosage of H2O2 and a shorter conditioning time. Contact angle measurements and X-ray photoelectron spectroscopy (XPS) analysis were employed to understand the depressing mechanism of these oxidation treatments. It was found that the chalcopyrite surface covered by various oxidation products that render the surface hydrophilic. Meanwhile, the surface of molybdenite was slightly oxidized, making the surface remained hydrophobic. INTRODUCTION Molybdenite (MoS2), the main source of molybdenum, is associated with copper sulfide minerals, mainly chalcopyrite (CuFeS2), as a trace mineral. Molybdenite is a valuable mineral and plays an important role in making the copper-molybdenum (Cu-Mo) processing plant economically viable (Laskowski, Castro, & Ramos, 2013). Therefore, it is important to recover molybdenite efficiently and effectively and separate it from copper minerals. Conventional separation of copper and molybdenum minerals is usually carried out in selective Cu- Mo flotation process using a copper depressant (i.e., sodium hydrosulfide (NaHS), Noke’s reagent, or cyanides) (Ansari & Pawlik, 2007; Bulatovic, 2007). However, this process still suffers due to imperfect molybdenum recovery. Moreover, the copper depressants use in this process are toxic and harmful to the environment. As alternative, various methods have been developed to replace these toxic and hazardous depressants, for example, by synthesizing new non-toxic or low-toxic depressants or by using various oxidation methods such as sodium hypochlorite in an alkaline medium, manganese dioxide, oxygen, and plasma pre-treatment (United States Patent Office Patent No. US 2559104 A, 1951; Patent No. US2238250 A, 1941; Barzyk, Malysa, & Pomianowski, 1981; Hirajima et al., 2014). However, the oxidation phenomenon and mechanisms are still not clear since detailed surface analyses have not done and the plasma pre-treatment method is facing technical issues for industrial scale application.

Jan 1, 2019