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By R. Ricketts, D. G. Vanderveer, F. Kirby

"A program of mapping, sampling and testing to determine the potential of the aggregate resources throughout Newfoundland and Labrador was commenced in 1978. A 6-km-wide corridor parallel to all transportation routes was selected as the study area.The study consisted of a preliminary surficial interpretation followed by field mapping and sampling. Granular resources were sampled at a 500-m interval and non-granular materials were sampled at a 1.5- to 2-km interval. Data were gathered on landforms and stratigraphy as well as a field description of the sample and deposit types. Particle-size and lithologic analyses of the samples were also conducted.The locations of samples and deposits were mapped on 1:50,000 topographic base maps and summarized for publication at a 1:250,000 scale. Surficial geology maps were produced at a 1:50,000 scale. A computer program is being developed for the search and retrieval of field data.The first phase of the Inventory of Aggregate Resources program is nearing completion and to date has collected information on approximately 9,500 samples throughout Newfoundland and Labrador. This data will be used to provide f or the effective management of these vital non-renewable aggregate resources.The mineral aggregate industry in Newfoundland and Labrador is very competitive, but undercapitalized when compared to other provinces. The demand for aggregates is forecasted to increase to in excess of23 million tons per annum by the year 2002.In terms of over-all mineral production, the extraction of sand, gravel and stone has a significant impact when compared to other minerals in terms of value and tonnage produced.IntroductionThe province of Newfoundland has approximately 18,000 km of coastline. A large proportion of the population lives in coastal communities where, traditionally, beaches have been a major source of construction materials. Beaches have, therefore, often been subjected to overutilization, resulting in erosion of back beach areas or adjacent headlands."

Jan 1, 1982

By A. A. Nourbakhsha

The main application of refractories based on andalusite in ferrous metallurgy are in air heaters of blast furnaces, mobile mixers, and hot-metal transfer, pig iron ladle, steel-casting, and intermediate ladles. Three specific properties of andalusite are very interesting for the production of high-quality shaped and unshaped mullite refractories: being free of water, its crystal a remains stable without changing its density or volume after firing and therefore has no internal porosity .Graphite additives due to its great wetting ability play an important role in raising the property of refractory bodies. In order to improve properties of andalusite graphite bodies, oxidant agent for increasing in hot strength and oxidation resistance can not be avoided. In this research, firstly 8 pig iron ladles from ESC(esfahan Steel company) with different conditions selected and monitored during usage to determine of main factors affecting on corrosion of refractories. Results showed that tap to tap time, undischarged pig iron ,penetrated pig iron are the main ones.Literature studies showed that the graphite addition help to thermal shock resistance and liquid penetration(steel,pig iron) effect. In this research particle size distribution of aggregates containing andalusite and graphite determined by andreazen equation. Then different mixes with different percentage of graphite,Si, FeSi, SiC prepared by uniaxially pressing under 1500kgf/cm2 pressure by adding 4%wt novalak resin and 0.4%wt HMTA as hardner. Samples fired at 1100-1500oC under reduction atmosphere. According to specific standards density, porosity and CCS measured. Analytical tests such as XRD, SEM were carried out on sample selected from point of view of mechanical properties at different temperatures. Results showed that optimized amount of Si and Fe-Si for _-SiC formation in body is 4 and 6 %wt respectively at 1500OC. XRD and SEM analysis on sample with Fe-Si additives showed no destructive phase through the matrix, also SiC additives not only increase oxidation resistance but also no effect on thermo-mechanical properties.

Oct 1, 2011

By C. R. Edwards

"Equity Silver Mines Limited owns and operates an open pit silver 111ine located 37 km southeast of Houston in the Omineca District of British Columbia, Figure l. The property is approximately 300 km west of Prince George and 417 km east of the Port of Prince Rupert. This is the largest silver mine in Canada and one of the five largest silver producers in the western world.Equity is a subsidiary of Placer Development Limited, which owns 70 of Equity's issued and outstanding Common Shares. Placer was responsible for financing and developing the property and now provides general management and administrative services.All mining is by open pit methods, using three 1600 P&H shovels, eight 45 t Caterpillar and three 77 t Wabco trucks, three 40R rotary drills, two D8 dozers, one 9888 front-end loader and other ancillary equipment. We are now phasing out the Southern Tail pit and bringing the Main Zone into productior; see Figure 2.Comminution in the concentrator is achieved via primary, secondary, and tertiary crushers followed by rod and ball mills. The flotation circuit, with Southern Tail ore, produces a concentrate containing between 3 000 and 6 500 g of silver per tonne, 201 copper, and 25 g of gold per tonne. Running Main Zone ore, concentrate grades will decrease to about 4 500 g of silver per tonne, 17% copper, and 18 g of gold per tonne. Design milling capacity is 5 300 tpd, which is often exceeded.The flotation concentrate contains relatively high levels of antimony and arsenic. Because of these impurities Equity was unable to obtain economically acceptable long-term sales contracts for flotation concentrate. For this reason a leaching process was developed, and a plant designed, built and placed in operation. Due to the complex and unique technology involved, many start-up problems were encountered but by March 1983 the leach plant was considered ""fully operational."" This paper describes some of our progress and problems in bringing the Equity leach plant through its start-up period."

Jan 1, 1984

By E. G. Joe

"The continuing increase in costs for power and fuel are affecting significantly the cost for milling of Canadian ores. In order to ascertain the scope and source of energy consumption in Canadian mills for further energy conservation opportunities, CANMET, in cooperation with the Canadian Mineral Processors carried out a detailed survey on the metallic mineral processing plants. Sixty-seven mills representing almost 300 million tons of annual capacity participated in the survey. The results indicated that in general, copper and iron ore mills, representing the bulk of the tonnage, consumed relatively low levels of energy per ton. Nickel-copper, zinc-copper and lead-zinc ore mills were moderate in energy consumption, while uranium and some copper-lead-zinc mills reported much higher energy consumptions due primarily to the increased heat requirements of the ore pulp. While the plant heating costs varied somewhat with the location in Canada, it was apparent that the lowest heating costs were achieved by those built with proper insulation and energy conservation design. If the cost per million British Thermal Units is taken as $2.00, the mill energy costs generally ranged from 40¢/ton up to $1.20/ton.IntroductionEnergy consumption in milling has always been of significant interest to Canadian mineral processors who traditionally calculate their milling costs down to pennies per ton. Even the measure of an ore's grindability is expressed empirically by Bond in kilowatts hours per ton milled. Some of the mills located in isolated areas were obliged to generate their own electrical energy from fossil or petroleum-based fuel at a cost substantially higher than that obtainable from captive or public hydro-electric installations. These mills practiced energy conservation long before the advent of O.P.E.C. energy cost inflation. One example of this was the use of electric boilers to produce process and heating steam during off-peak electrical load periods."

Jan 1, 1978

Everybody would agree that a good roadway leads to more efficient haulages by reducing haul times, smoother and safer operation for operator and less wear and tear on tires and equipment, thus improving availability and overall costs efficiency. There is a number of factors that come into play to achieve this especially when developing a mine. Excavations of Ramps and Headings: Operators should follow surveyors and layouts accurately while driving ramps. To maintain constant grades on ramps, which leads to better design control of water and easier grading of roadways. Dips and dives lead to needless over excavations and unnecessary cost and disposal of waste rock. When dressing a ramp some areas may require 2 to 3 feet of material to correct the dips and perfect to constant grade. You may ask yourself, is this all necessary? A small exercise is to follow a loaded haulage truck up ramp. At the first hump, grade change, notice the down shift and notice how far the truck travels to regain the up shift and his speed. The difference is remarkably noticeable when the grade is constant. All this leads to less driveline shock, less horsepower demand, better fuel efficiency and more productivity. Drill & Blast Patterns to control water for Ramp Headings: The drill pattern needs to allow for a ditch. A slight cross fall to form a ditch can easily be done with the lifters. The preferred side of the ditch is always the inside curve of the ramp as it has less tendency to erode roadbed material. On a level, the ditch is dictated by the sump location. A blasted ditch will eliminate water pounding and keep the roadbed dry. Water trapped in roadbed is the major cause of rollout and de-stability of roadbed. If the water is allowed to drain in the middle of the ramp it will cause erosion, movement and loss of fines and an increase in grading intervals.

Feb 1, 2008

"The ore currently being milled at Delnite is a high pyrrhotite ore with the gold intimately associated with the pyrrhotite. To get maximum recovery of the gold with the lowest reagent cost, it is necessary to keep the pH as low as possible In the laboratory the best results were obtained with a pH of 10 but in the plant it was found that a pH of 11 was the minimum practical pH. With the n1anual control it was difficult to maintain this figure.The Mines Branch had a conductivity probe (see Fig. 1), developed by Dr. G .G. Eichholz that they wished to try out in plant operation.On June 17, 1963, the probe with all necessary control equipment was installed in the regrind circuit (see Fig. 2) at Delnite Mines. Crushed quick lime is fed at this point but it is also necessary to feed hydrated li1ne at two other points in the cyanide circuit in order to maintain the pH close to 11. 0 throughout the circuit. Only the feeding of the quick lime at the regrind mill was controlled in this test.One thing that concerned me with the operation of the probe was whether the reading on the probe chart would correspond to the pH. A look at the graph accompanying this paper will show that except for very minor variations the graphs of the chart reading and the pH reading are parallel. The upper graph shows that the pH reading and the lime titrations also correspond and that the cyanide titrations appear to have little effect on the pH reading.Table 1 gives the results obtained for the two months before and after the instal1ation of automatic control as well as the two months of control. Unfortunately there was a decided change in ore just after the automatic controls were removed. There was much less pyrrhotitc in the ore than usual thus changing the need for lime and cyanide."

Jan 1, 1964

NANA Accepts Mining ­ 1974 - Leadership presented mining option to people ­ Community meetings; recognized need for cash economy ­ 1979 - Poll shows consensus in favor of mine development ­ 1980 - Shareholders approve mining; Congress approved land selections

May 1, 2002

By D. K. Storer, W. Lee, S. Tkach

"A microcomputer-controlled multi-probe K-40 monitoring system was developed internally for the Phase II mill at the Lanigan potash mine. The K-40 probes were mounted in a variety of positions in several process streams, ranging from dry ore on a conveyor belt to brines and slurries. A description will be presented of the objectives, the equipment and the difficulties encountered in correlating results.IntroductionThe successful control and optimization of most metallurgical processes involves monitoring not only physical parameters such as flow rate and temperature, but also the chemical compositions of critical streams. Composition has proven difficult to measure on -line in industrial scale processes. However, in potash processing there is a convenient chemical parameter that has long been utilized to measure in situ the concentration of potassium chloride (KCl). Potassium has a long-lived isotope which is slightly radioactive. K40 decays by beta elimination to Ca40, ultimately radiating a characteristic gamma-ray photon. This gamma emission when detected by scintillation counter can be used to estimate the concentration of potassium in a process stream."

Jan 1, 1990

Overview of Presentation ?About MiHR ?Human Resources Challenges in the Minerals and Metals Industry ?Getting the Facts -Sector Study Project ?Key Findings ?Recommendations ?Conclusions ?MiHR?srole

May 1, 2006

"Cambior now holds a 49% interest in the Valdez Creek Placer, having acquired American Barrick's 23.125% interest in May 1990. At the same time, the company assumed the management of the operation and received environmental permits to divert the river. Plans are underway to divert the Valdez River in order to continue mining the deposit upstream, and to construct a new, higher capacity wash plant (by 60%) closer to the operations. Mining operations, which were suspended in October 1989 pending higher gold prices, resumed in August 1990. The mine is expected to produce 10000 ounces of gold in 1990and more than 80 000 ounces in 1991.IntroductionThe Valdez Creek Placer is located in the Valdez Creek mining district of Alaska on the western flank of the Clearwater mountain range, about 150 miles northeast of Anchorage by air (Fig. I) . The mine is accessible by air or by road.The mine is situated at an elevation of about 3000 ft and is isolated from towns and villages. Consequently, the mine site contains the facilities necessary to house and feed 135 people at any given time. The Valdez Creek Placer has 159 employees, including 130hourly and 29 salaried employees. The hourly employees are divided into three rotating teams. The daily schedule consists of two shifts of 12 hours, and each team works 14 days followed by seven days off, resulting in a weekly rotation of teams.The salaried employees work normal hours for ten consecutive days followed by four days off. A weekly rotation ensures continuity of services and management.Gold was first discovered at Valdez Creek in 1903, when a team of four men led by Peter Monahan left the village of Valdez to explore the area up the Susitna River, to the east of Mount McKinley. They discovered gold at the mouth of a creek that they immediately named Valdez Creek, in honour of their native town. In 1904, they discovered the Tammany Channel to the right of Valdez Creek. This was the beginning of a long period of mining in the area which continued until the beginning of World War II .In 1980, the Denali Mining Co. was formed. Geophysical work and drilling carried out through 1980 and 1981 led to the discovery of Channel A, a second channel which is older than the Tammany Channel. In 1983, Denali Mining optioned its claims to Camindex to explore, develop and mine the Valdez Creek property. A joint venture agreement was signed among Camindex, Talcorp, BCP Alaska (Barrick), Sum Resources (Cambior), Norfran and WGM to form Valdez Creek Mining Co., Inc., and production began in 1984. Valdez Creek Mining Co., Inc. (VCMC) was subsequently restructured and in 1989 was held by the following three companies: Camindex Mines Ltd. - 51070; Cambior Inc. - 25.875%; and American Barrick - 23.125%.VCMC is managed by a committee of the partners. Decisions made by the committee are subject to a 70% majority vote. Cambior receives its share of production in kind, and contributes in the same proportion to the expenses of the joint venture (VCMC)."

Jan 1, 1990

By R. L. Eadie, J. L. Luo

The effect of plastic deformation on near-neutral pH stress corrosion cracking (SCC) also known as environmentally assisted cracking was studied by cracking rate measurements. All the tests were conducted in a synthetic near-neutral pH solution. The traditional potential drop tests using compact toughness (C-T) specimens with 2-3 mm long pre-cracks showed varying plastic deformation by rolling would increase the material's susceptibility to SCC. This was confirmed by cyclic loading tests using double notched flat tensile specimens with 200-300 µm long pre-cracks which gave the same results that the specimen with large scale plastic deformation by tensile loading had a relatively faster crack growth rate than that with only small scale plastic deformation.

Jan 1, 2005

By Forrest A. Kerr

Introduction During recent years the work of the Geological Survey in northwestern British Columbia has been in the nature of reconnaissance and exploratory mapping, largely for the purpose of gaining knowledge of the potential mineral-producing zones known to occur along the eastern side of the Coast range, and also to gather such data as might facilitate their exploitation. Commencing with fairly detailed mapping in the more accessible parts, the work has been continued and expanded until now an understanding has been gained of the more general features of a continuous section along the range for a distance of about two hundred miles. Geographically, northern British Columbia is more or less isolated. It is cut off from the central part of the Province by a rugged and mountainous country which, except for the telegraph line and a rarely travelled trail along it, is not penetrated by any developed routes or lines of communication. To the west, it is cut off from the Pacific by the Coast range, an extremely rugged area sixty to eighty miles in width which, except along the few navigable rivers cutting across it, cannot be readily traversed. Politically, it is further isolated in this direction by the narrow strip of south-eastern Alaska which borders the coast and extends inland to the axis of the range.

Jan 1, 1931

By Bennett McLaughlin

The Penna Shaft is the deepest single?lift shaft in North America. AGNICO-EAGLE started sinking the Penna shaft in 1995 with a hoisting plant capable of hoisting 1 800 tonnes per day from a 2 240-metre deep shaft. Exploration success at depth lead to the replacement of the existing double drum hoist by another double drum hoist capable of hoisting 7 000 tonnes per day. The use of another double drum hoist, instead of a Blair Multi-Rope winder, to hoist 24,5-tonne skips was only made possible by lowering the safety factor. This paper describes recent experiences at Penna and in South Africa, and examines the possibility of extending the double drum hoisting system to a depth of three kilometres.

May 1, 2003

By F. W. Wellmer

"This paper discusses the changing nature of reserve/consumption ratios, using data from Canadian operations and focussing particularly on copper.IntroductionStudies in world metal consumption and world metal reserves are done regularly by the U.S. Bureau of Mines, the German Federal Institute of Geosciences and Raw Materials, and other institutions. Formerly, on the basis of these data, the following figures were calculated:• a static lifetime of reserves-reserves at year end divided by the consumption of the previous year;• a semidynamic lifetime of reserves-reserves at year assumed constant growth rate.These figures were severely criticized because they were purely theoretical and did not consider the availability of the raw materials.Recent practice is to calculate reserve/consumption ratios (R/C ratios) for periods(1) or for different years(2,3) and to interpret the trend. For base metals, these R/C ratios are increasing (Table 1). This had led to the conclusion that today the relative availability of base metal reserves has increased as compared to 10 or 15 years ago(4,5) As there are normally numerous deposits for each commodity, the anticipated production rates of deposits are irrelevant to this conclusion as long as the statistical distribution of possible production rates does not change significantly.The above conclusion is only justified if the quality and the nature of the data base remain the same. This is certainly the case for the consumption data. It is also assumed for the quality of the reserve figures, which usually take into account only proven and probable reserves. For the nature of the reserve figures, however, this does not seem to be true.Undoubtedly, the reserves increased in absolute terms between 1966 and 1978. To put the absolute reserve figures into perspective, it is common, as said above, to compare them with the mine production, which increased in the free world between 1966 and 1978; e.g. for copper by 40%, for Pb by 17% and for zinc by 36%. However, the reserves are also influenced by other factors, such as average size of the mines, changes in the grade of the mined ore, price-to-cost relationships and the like.The purpose of this paper is to examine the nature of the reserve figures in more detail. Because the completeness of mining data published in Canada is rather exceptional, Canadian .data serve as an example and it is assumed that the conclusions can be extrapolated to other mining countries or at least to those with free market systems."

Jan 1, 1981

By Tong Qin, Jinxia Zhang, Fusheng Niu, Limei Bai

"The interaction between a deformable bubble and a rigid particle in viscous liquids under microgravity conditions is studied by direct numerical simulation via an Arbitrary-Lagrangian-Eulerian (ALE) method. Due to the microgravity, the bubble will stay as a sphere in the fluid without external forces and its deformation is caused by the moving particle. In this study we assume the flow field is axisymmetric and the motion of the gas inside the bubble is neglected. The particle equations and the Navier-Stokes equations for the fluid are solved in a unified finite element framework. The particle obtains an impact velocity by an external force and the impact velocity is varied over a wide range. A ‘critical time scale’ T0 was identified for the interaction process. For successful contacts between particle and bubble, the time scale of their relative motion must be less than T0.INTRODUCTIONThe particle-bubble interaction is an important process in many applied industrial domains such as water treatment, froth flotation and liquid steel purification. It is a complex three phase process in that it embraces many research areas: fluid hydrodynamics, interfacial forces, particle and bubble behaviour and solution chemistry.Derjaguin and Dukhin(1961) clarified the particle-bubble interaction by introducing three-zone model for small and medium size particles. When the particle is far away from the bubble surface, the hydrodynamic forces are dominant. The particle is driven toward the bubble surface by the particle inertial and gravity. This region is regarded as Zone 1 and the interaction in this zone is a collision subprocess. As the particle gets close to the bubble, it exits a zone where the liquid flow around the bubble surface creates a tangential stream from the upper surface of the bubble toward the rear of the bubble. This is Zone 2 which is called the diffusiophoretic zone. In Zone 3, the particle is very closed to the bubble, surface forces like van der Waals, electrostatic and structural forces are considered. Contrast to Zone 1, Zone3 can be regarded as an attachment subprocess."

Jan 1, 2016

By F. Goodarzi

"The nature and distribution of the inorganic components and the concentration of certain environmentally sensitive elements, such as, As, Sb, B, Hr, Cd, CI, F, Mo, Pb, S, and Se, or elements detrimental to the utilization by industry such as B, Na, P and S are very important in determining coal quality. The concentration of elements in coal is controlled by the following factors: (1) frequency of occurrence of partings; (2) nature and habit of occurrence of minerals in coal; (3) environment of coal deposition; (4)rank; (5) geological setting; and (6) weathering.Selective mining of coal can reduce the input of undesirable elements such as As and S by eliminating the part of the seam which is enriched in these elements. IntroductionCoal is a sedimentary rock of plant origin and consists of two components: an inorganic component (minerals) which is formed by influx of clastic sediments into the coal-forming environment and also by activities of groundwater, and an organic component (macerals) which is formed from plant debris deposited in the original coal-forming environment. However, coal is a generic term with a broad range of utilization characteristics. A coal that is attractive for use in one market - such as generation of electricity (combustion) - may not necessarily be acceptable for use in another, such as for production of coke (Davis, 1978).The elements in coal are associated with both macerals and minerals. The elements associated with coal macerals, such as C, H and N are termed elements with organic affInities. The concentration of elements with organic affmities (including B, Br, CI, F, I, S, and Mg) often decrease with increasing ash content of coal (Nicholls, 1968; and Finkehnan,1982). FIgure 1 shows such a trend for chlorine. Some of these elements may also be incorporated in coal maceral as very fine mineral grains (Finkelman, 1982). The elements with organic affmities are often volatilized during combustion and carbonization of coal and emitted in the vapour phase. Emission of some of these elements (C, 0, Br, N and S) is of major environmental concern (Lim, 1979)."

Jan 1, 1994

By David Minson

On many occasions, operators have seen that high density and paste thickener performance does not meet the initial design expectations, and the installed equipment produces lower than expected thickener underflow solids content. This lower-density tailings slurry leads to more water and lower beaching angles being seen at the tailings storage facility (TSF). This not only reduces storage capacity, but also increases the risks of non-compliance with environmental permit constraints. Reasons for the shortfall in performance can be many, but perhaps the main problem is inadequate understanding of the critical components of thickener design and operation, with too much reliance placed on standardised design terminology; such as tph/m2. This paper discusses additional factors that influence high-density thickener operation and presents test work that defines “Dynamic Slurry Fluidity”, “Bed Compaction Rates”, and “Natural Slump Angles” for these high yield stress/viscosity slurries. Finally, the paper uses this data to define thickener geometry, drive torque, floor slope(s), rake blade configuration and how this can be used to provide engineers with more tools for in depth evaluation of a High-Density thickener design. INTRODUCTION Generally copper mines are high tonnage operations, with copper flotation concentrators producing high volumes of tailings slurries of milled solids and water. As a result, Tailings Storage Facilities (TSF) become very large and this, along with potential scarcity of water, favours a design that maximises storage space. In theory, the use of filtration equipment that achieves immediate water recovery, high TSF in-situ densities and high stacking angles are preferred for this duty. However, large tailings filter units are uncommon, and have high operating and capital costs that favour the use of thickeners. The use of high-density thickeners to produce a thick or near-paste like slurry with high un- sheared yield stress values above 100 Pascals, is becoming more common. The higher thickener underflow is required to not only maximise water recirculation within the mill, but to also increase beaching angles, maximise tailings storage capacity and minimise evaporation losses arising from large areas of superficial water. However, the reliability of this equipment is also under scrutiny and there is perhaps a need to better define these designs.

Jan 1, 2019

By BARRY BANNATYNE

Gypsum in Manitoba is quarried from the Jurassic Amaranth Formation a1 Gypsumville and Harcus; it has been mined near Amaranth and Silver Plains. At Gypsumville, gypsum is carbonate- free but has thin clay laminae between gypsum layers; the rock is folded in places, and some anhydrite is present. At Harcus, gypsum contains scattered dolomite, both as patches in a penemosaic pattern and as thin layers; small chalcedonic concretions also can be found. The quarry at Harcus was opened late in 1978 and currently produces about 100,000 tonnes annually. The two inactive mines near Amaranth operated from 1932 lo 1963 and from 1967 to 1970, and contain a central anhydrite layer. The mine near Silver Plains was flooded from an underlying artesian aquifer in 1975 and has not been reopened. Manitoba gypsum is used in cemem plants in Mani1oba and Saskatchewan and in wallboard plants in the three Prairie Provinces.

Jan 1, 1984

By Yuki Soma, Tsutomu Ogawa, Jae Hyung Hong, Toshihiro Nagato

The Saganoseki flash furnace was built in 1973. The steel frames of the flash furnace were designed to absorb the expansion of refractories and contain expansion-absorbing materials between their frames and the skin plate. Expansion and deformation resulted in the furnace settler because of the repeated increase and decrease in temperature from more than 30 maintenance shutdowns. Preliminary investigations were conducted to confirm the condition of the settler, and it was found that partial cracks accompanied the expansion in the lower layer of the refractories. As a result, the risk of molten-matte leakage increased. Based on the above investigation, it was decided that the entire settler should be replaced by introducing a spring-displacement structure that was capable of maintaining the furnace integrity over an extended period. The difficult redesign and rebuilding of the Saganoseki flash furnace was accomplished successfully and safely in 76 days from September to December 2017, as planned. Because of various redesigns, the furnace was changed to a flash-smelting furnace that can withstand higher thermal loads based on recent trends of copper concentrate with increased sulfur and reduced copper. The new furnace is operating smoothly at present. INTRODUCTION The flash-smelting furnace of Pan Pacific Copper Co., Ltd. Saganoseki Smelting & Refinary commenced operation in 1973, with an original designed feeding rate of 71 t/h. As a result of facility remodeling and operational improvements, shown in Figure 1, the feeding rate was increased to 215 t/h in recent years. After more than 40 years of operation, the original settler in the melt holding section of the flash furnace underwent significant expansion and deformation.

Jan 1, 2019

By S. Malkin, Z. Shi, H. Attia, D. Chellan

An experimental investigation is reported on creep-feed grinding of a nickelbased alloy (IN718) using conventional aluminum oxide grinding wheels. Straight surface grinding expe.riments were conducted on nickel-based alloy blocks over a wide range of grinding conditions with straight oil as- grinding fluid applied at high pressure. Grinding power, forces, workpiece surface roughness, and radial wheel wear were measured. Specific material removal rate as high as 64 mm2/s was Used without burning the workpiece, which can be attributed to the extremely low specific energy of 12.6 J/mm3. It was found that the grinding wheel under the present test conditions had a strong capability of self-sharpening as indicated by the stabilized power, forces, and surface roughness after an initial short transient.

Jan 1, 2006