Search Documents

By J. Walenzyk

This report investigates the limited use of super large backhoes in surface mines around the world. The demand for super large (> 700 t) hydraulic excavators is rising and most of the machines are sold as front shovel (FS) application. Despite advantages in faster cycle times super large backhoe (BH) machines are a minority (ca. 1% of the total machine population). This investigation is based upon the comparison between the volumes available to the BH and FS in one block and before the machine needs to move. The working conditions are created with standard parameters as given by Original Equipment Manufacturers of hydraulic excavators regarding bench height, angle, and material density. The travel amount of both machine types is derived from these volumes and compared. Furthermore, the travel amount of super large BHs is compared to the travel amount of smaller mining BHs and evaluated based on attachment length and bucket size. The calculations conclude that the volume of material recoverable in a single spot is about 75% higher for FSs. This leads to an increased travel for BHs. The combination of both the increased appearance of electrical driven machines (in the class of > 700 t excavators) and the travel amount lead to a high risk of damaging the cable or the excavator. Enlarged wear on the travelling parts is also attended by the higher travelling. A direct comparison of commonly found working set ups shows that the ratio between volumes is 32% which leads to an increased travel amount. In comparison with a 250 t BH, the calculation shows that the smaller BH is able to reach nearly 50% of the volume of the super large BHs. The main reason for this is the respectively high reach on smaller machines which, combined with other factors, leads to a significantly higher number of BHs in lighter classes.

Aug 1, 2013

By R. O. Burt, R. W. Bruce

"IntroductionGravity concentration, the separation of two or more minerals due to the difference in their relative specific gravity and usually in a fluid medium, is the oldest recorded form of mineral processing.The principles of gravity concentration were known to, and practiced by, the ancients and Hoover's commentary to De Re Meta/lied'' notes many early references.Over the centuries, as direct smelting ores became more and more uncommon, the mineral processing industry developed and along with it many varied methods of concentration. Gravity concentration declined rapidly in importance in the early twentieth century, and it is only in the last decade that its advantages are being re-evaluated.A significant proportion of ores are processed by gravity; a large proportion being in the iron and coal industries<'>, or in the treatment at very high volume of relatively coarse alluvial deposits. The tin dredges of South East Asia and Beach sand plants of Australia are prime examples.An indication of the broad application of gravity concentration is the variety of minerals presently recovered. A list is given in Table 1; this list is not exhaustive.Principle of Gravity ConcentrationGravity concentration methods separate minerals of different specific gravity by their relative movement in response to gravity and one or more other forces, the latter often being the resistance to motion offered by a viscous fluid, such as water or air.The relative movement of minerals to one another is achieved by a variety of methods. These are: separation in a fluid of intermediate density (heavy media separation); by stratification (such as in jigging); in a flowing fluid (cones, spirals, sluices etc.) and with a horizontal shear (shaking tables etc.)."

Jan 1, 1989

By Franc R. Joubin, Donald H. James

"THE PROPERTY of Rexspar Uranium and Metals Mining Co. Limited is situated within 3 miles of Birch Island, British Columbia, a village on the C.N.R. main line 80 miles north of Kamloops. The claims extend up the south slope of the North Thompson River valley.Prospecting in the area dates back many years. Fluorite was reported in 1918, lead showings in 1926, and bog manganese in 1929. Diamond drilling in 1942-43 indicated more than a million tons of material containing 20 per cent fluorite and 15 per cent celestite in what is known as the 'fluorite zone'. Marketing and beneficiation difficulties have delayed further development of this deposit.Systematic checking of specimens and drill core by the B.C. Department of Mines and the Geological Survey of Canada revealed low radio-activity in the fluorite zone material. Subsequent prospecting with Geiger counters revealed zones of higher radioactivity a few hundred feet distant from the fluorite zone. The property at this stage came under the development direction of Technical Mine Consultants Limited, who undertook a program of diamond drilling. Of the several radioactive zones located, two have so far rewarded development and appear to constitute low-grade uranium orebodies. These are referred to as the ""Black Diamond"" or ""BD"", and the ""A"" zones -see section. Open-pit mining of the thick, dip-slope orebodies appears possible.General GeologyWalker (1) describes the local rocks as metamorphosed sediments cut by granodioritic intrusives and by greens tone. At the property the sequence is:GreenstoneUnconformity or intrusive contact Trachytic member -tuffs, breccia, flows. Argillite -black, thin, not always present. Schist -thin-bedded sericitic quartzite and siltstone with some calcareous layers."

Jan 1, 1949

"Line Creek Resources Ltd. produces three million clean tonnes per year of metallurgical and thermal coals for the international market. The mining operation, which is located in the mountains of southeastern British Columbia, produces coal from four pits by conventional truck and shovel methods.The South Pit, which has been in operation since J982, has a benched highwall 490 m in height. Shallow dipping coal seams at the base of the syncline formation will be exposed in the highwall over the next four years of pit life. The first of these seams, which will measure 500 m in final highwall length, was exposed in J994. An auger, powered by a 1250 hp engine capable of cutting 2.2 m diameter holes, was brought on site in J994 and has proven to be a successful means of recovering coal from the J3 m thick seam.This paper covers the background to the selection of an auger mining method to recover highwall reserves together with the geotechnical and methane considerations for this approach.IntroductionLine Creek Resources Ltd., a wholly owned subsidiary of Manalta Coal Ltd., owns and operates the Line Creek mine. The operation which consists of four active pits, a processing plant, and rail loadout is located in the southeastern coal fields of British Columbia. Figure 1 shows the general area.The coal bearing sequence at the mine is preserved on the east and west limbs of a major syncline. Up to twelve mineable coal seams, ranging in thickness from 2 m to 15 m are contained in the formation. Within the pits the seams are stratigraphically consistent and structurally undisturbed, except for small-scale thrust faults that create multiple repeats of the seams."

Jan 1, 1995

By R. W. Pugh, S. L. Ma

"High carbon rods are typically used as the grinding media in rod mills. Usually an AISI I 090 grade of steel is used which provides the high hardness needed for abrasion resistance and sufficient toughness to prevent rod breakage.Work to improve performance in the past has mainly concentrated on increasing the hardness for improved abrasion resistance. This has been accomplished by adding alloys to the steel which uniformly increase the hardness across the rod section.However uniformly increasing the hardness of the rod produces a more brittle product and at high hardness l1evels can lead to rod breakage problems. Heat treatment processes can be used to selectively increase the surface hardness to high levels while maintaining the core at lower hardness levels to give a product that wears slower without breakage. AltaSteel has developed and mill tested a higher hardness heat treated grinding rodThis paper describes the development of a simple model to predict performance improvement with different hardness profiles and the AltaSteel heat-treated grinding rod process. Mill characteristics of the Endako operation are summarized and predicted performance improvements based on hardmess are compared to the approximate 25% achieved in trials running since July, 1991 at this mill. Early results of other mill trials are also presented.INTRODUCTIONAltaSteel, in Edmonton is a major supplier of grinding media. We have a strong desire to continually increase the quality and performance of our products and thereby improve the efficiency of our customers' processes. In the area of grinding rods, we wanted to produce a product with a significant improvement in wear rate. Second to energy, rod consumption usually represents the next highest cost in processing. This paper presents our development of a heattreated grinding rod that has reduced consumption by about 25% in trials to date."

Jan 1, 1993

By S. Demetrio

China has a traditional segment of Copper Smelters with its own Chinese technology, typically small in size from 20.000 to 50.000 tCu/a, and they are responsible for the processing and production of more than 600,000 tonnes of blister copper per year. For occidental experts, like one of the authors, it is not so easy to believe that they are profitable, because the size and the use of old Blast Furnaces, but they are with simple and effective production lines. The authors are convinced that China has a good alternative to improve theses old copper production lines with small innovative changes in its process arrangement, jumping to higher level technology, more efficient in production, productivity and cost, assuring copper as a raw material for the country and its high speed development. The key to improve that the traditional Chinese Copper Smelter is white metal as an extra raw material as a first step doubling the copper production from raw materials and the complete replacement of concentrates by white metal (WM) in a second step increasing it four times or more as the final goal. It is not so difficult nor expensive to make it happen, but there are a couple of questions for implementing this innovation in the copper smelter industry: 1 - Will the old fashioned mentality be open to this new direction for China being the biggest copper consumer on one side, and the biggest copper Primary Producers on the other hand? and 2 - Will they be open to follow this "win-win" strategic way for both ends of the copper market? We don&apos;t know if anybody will take up the challenge to improve it but our responsibility as a consulting expert and a university professor is to put the alternative on table for others to decide whether to implement it.

Jan 1, 2007

By James F. Allan

"The Kipawa yttrium-zirconium prospect is located about 150 km south of Noranda, Quebec. Yttrium-zirconium mineralization occurs in an elongate, regionally metamorphosed and deformed peralkaline complex in the Grenville terrain. The mineralized zone, as defined by =.0.10% Y2O3, is 1300 m long and 10 m to 30 m wide. Zirconium content typically ranges from 0.50 to 1.20% ZrO2, The mineralized zone strikes northwest and dips 20 degrees to 35 degrees to the southwest. It occurs at the structural base of the peralkaline complex in syenite gneiss and calc-silicate rocks and is conformable with both the footwall contact of the complex, and the layering and gneissosity in the enclosing rocks. Minerals of economic interest are eudialyte (sodium-zirconium silicate), mosandrite (sodium-titanium silicate), britholite (calcium-silica phosphate), and a few other uncommon zirconium and titanium silicate minerals.The highest and most continuous yttrium grades occur in the upper part of the syenite gneiss unit in areas containing eudialyte, mosandrite and minor britholite. Yttrium also occurs in britholite in calc-silicate rocks and in mosandrite in syenitic rocks in the lower part of the syenite gneiss unit. Both the yttrium-bearing minerals and yttrium grades are erratic in the lower part of the unit. Zirconium is generally more uniformly distributed than yttrium. A peralkaline granite gneiss phase of the complex intrudes the syenite gneiss and possibly truncates it, and the mineralization, near the southern boundary of the mineralized zone.The mineralized zone is coincident with a radiometric anomaly and discontinuous yttrium and zirconium soil anomalies.More drilling and metallurgical studies will be required to determine the economic viability of the yttrium and zirconium mineralization at Kipawa."

Jan 1, 1992

By James F. Allan

"The Kipawa yttrium-zirconium prospect is located about 150 km south of Noranda, Quebec. Yttrium-zirconium mineralization occurs in an elongate, regionally metamorphosed and deformed peralkaline complex in the Grenville terrain. The mineralized zone, as defined by =.0.10% Y2O3, is 1300 m long and 10 m to 30 m wide. Zirconium content typically ranges from 0.50 to 1.20% ZrO2, The mineralized zone strikes northwest and dips 20 degrees to 35 degrees to the southwest. It occurs at the structural base of the peralkaline complex in syenite gneiss and calc-silicate rocks and is conformable with both the footwall contact of the complex, and the layering and gneissosity in the enclosing rocks. Minerals of economic interest are eudialyte (sodium-zirconium silicate), mosandrite (sodium-titanium silicate), britholite (calcium-silica phosphate), and a few other uncommon zirconium and titanium silicate minerals.The highest and most continuous yttrium grades occur in the upper part of the syenite gneiss unit in areas containing eudialyte, mosandrite and minor britholite. Yttrium also occurs in britholite in calc-silicate rocks and in mosandrite in syenitic rocks in the lower part of the syenite gneiss unit. Both the yttrium-bearing minerals and yttrium grades are erratic in the lower part of the unit. Zirconium is generally more uniformly distributed than yttrium. A peralkaline granite gneiss phase of the complex intrudes the syenite gneiss and possibly truncates it, and the mineralization, near the southern boundary of the mineralized zone.The mineralized zone is coincident with a radiometric anomaly and discontinuous yttrium and zirconium soil anomalies.More drilling and metallurgical studies will be required to determine the economic viability of the yttrium and zirconium mineralization at Kipawa."

Jan 1, 1992

By V. Caps

Gold has been often used as an alloying component in order to increase or modify the catalytic properties of an active metal. For example, in the hydroconversion of methylcyclopentane, Riahi et al [1] shown that the addition of gold to palladium or platinum enhances the activity which becomes greater than that of the pure metals whereas the selectivity towards products resulting from ring expansion increases. Molenbroek et al [2] presented a detailed study on Ni-Au catalysts which exhibit activity for steam reforming and have been proved to be more resistant toward carbon formation than the pure Ni catalysts. Blocking of highly reactive Ni edge and kink sites by Au atoms is assumed to be the reason for the increased robustness of the Ni-Au catalyst. Nevertheless, the own catalytic properties of pure gold have been almost ignored up to the pioneering work of Haruta et al [3] which demonstrates the high activity of gold for CO oxidation at low temperatures. Other reactions such as WGS reaction, propylene epoxidation or benzene total oxidation have then been shown to be catalyzed by gold [4]. This discovery induced the publication of papers devoted to analyze the different factors controlling the activity of gold catalysts. Among them the most relevant appears to be the gold particle size, the nature of the support [5] and the preparation methods [6]. Clearly distinguish between support and size effect is not straightforward. Indeed, the calcination step often needed to synthesize supported catalysts on different supports, leads to different particle sizes showing that particle size and support are interdependent. To circumvent this difficulty, Au-based supported catalysts have been synthesized by using laser vaporization and low energy cluster beam deposition. The as prepared catalysts have then been characterized and tested in CO oxidation reaction.

Oct 1, 2003

By V. Ogarlaci

Over 90% of the Romania&apos;s lignite reserve is located in the region of Oltenia, which is why this area is given special attention in the country&apos;s coal resources. Because the lignite deposit of Oltenia is comprised of 1-8 exploitable coal layers, their capitalizing calls for urgent arrangements to ensure rational exploitation under conditions of total safety, minimum losses and efficiency. The technology used the coal pits of North-Western Oltenia is the technology of excavation, transport and dumping in a continuous flow by using the complexes of excavation, transport and high-capacity dumping. The theme of this paper is of particular importance because the rhythmic operation in complete safety, is necessary both for sizing and keeping within the bounds of acceptable values, throughout the process of excavation of the geometric elements of the coal faces, marked by: the height of benches, slope angle and width of berms. Geometric dimensions of the benches of the open pits, the height of benches, slope angle and width of berms shall be calculated and shall be based on: the massif structure that forms the benches; physical-mechanical characteristics of rocks from the bench; working conditions, characterized mainly by mining and extraction methods applied; functional parameters of the mechanical equipment used to perform various operations on the coal face etc. For this study it was necessary a broad field for task tracking, measurement, research and design elements of the coal face, for land survey and a desk activity by: processing of topographic data, laboratory analyses on physical and mechanical properties of rocks and numerous specialized studies. In the final part, we bring some contributions to elucidating the issues of special importance for the lignite coal pits of Oltenia, coming up with data, graphics, formulas, solutions and concrete dimensioning of coal faces in terms of using technology in a continuous flow.

Aug 1, 2013

By S. Komar Kawatra, Youlian Zhou, Joseph A. Halt

"Bentonite clay has been the most common binder for iron ore pellets. Although bentonite increases pellet strength and durability, it has several disadvantages. For example, its use increases the cost of pellets and decreases pellet purity due to its composition and the higher levels of flux required to meet basicity limits. Several new types of binders have been proposed to replace clay. The binders include starch, modified humic acid (MHA), and dispersants. These binders were used to pelletize a magnetite concentrate from the United States. Typical doses of each binder type were 0.05-0.1 % starch, 0.5-0.75 % MHA, and 0.05 % dispersant. At these doses, the pellets exceeded the minimum recommended values for wet drop, dry compression strength, and fired compression strength. Pellets made with the MHA and dispersants had the lowest abrasion rates of the new binders, and were comparable to pellets made with bentonite clay. In this paper, we compare the new binders, considering pellet quality and the advantages and disadvantages of each binder type.INTRODUCTIONIron ore pelletizers need binders in order to make good quality iron ore pellets. The typical binder is bentonite clay, an alumino-silicate material. Bentonite increases the ballability and strength of iron ore concentrates, but reduces pellet iron grade and reducibility. As high quality bentonites can be expensive to ship, low quality bentonites are often used at significantly higher doses – this reduces iron grade and reducibility even further. Thus, several new types of binders have been proposed to replace bentonite clay.In this paper, we present several new binders for iron ore pellets: starch and modified humic acid (MHA). These have been tested alone and in mixtures with dispersants, or else with bentonite clay. Starch, MHA and their mixtures are briefly described below."

Jan 1, 2016

By Y. G. Seredkin, A. A. Vasiliev, A. Y. Senchenko, A. V. Aksenov

"Autoclave oxidation, bacterial and chemical liberation, roasting, acid oxygen processes, atmospheric oxidation prior to cyanidation are applied to process refractory sulphide concentrates. It is obvious that the processes with the maximum effect regarding gold recovery are those in which complete destruction of sulphides takes place and gold becomes accessible for further cyanidation through liberation from sulphide “capsula”. But the significant recovery increase is often accompanied by a process complexity, high expenses for equipment purchase and maintenance. It makes material processing costinefficient. Process optimization has led to progress in ultrafine grinding, bacterial and autoclave oxidation processes. In our opinion sulphide atmospheric oxidation processes are the most interesting as oxidation there runs without high temperature (up to 100°?) and excess pressure (open units), and with acid environment close to neutral. The process of sulphide concentrates treatment named “Albion” (Albion Process™) represents a combination of ultrafine grinding and oxygen oxidation without pressure and external heating. The process has a number of advantages: sodium cyanide consumption decrease during gold leaching and high ecological compatibility. TOMS institute in collaboration with Glencore Technology Pty ltd (Australia) tested Albion technology aiming at liberation of finely disseminated gold from flotation concentrate produced from refractory ore of one Russian deposit. The testing has shown that Albion technology makes it possible to process refractory products with quite small capital investments. It enables involving medium and small refractory gold deposits into processing.INTRODUCTIONThe considerable part of the world Au-bearing reserves is present by refractory ores with finely disseminated gold. Gold in such ores is often associated with sulphide minerals and locked in so called “sulphide capsula” since it is closed inside dense sulphide grains and inaccessible to recovery by cyanidation. To treat refractory sulphide concentrates containing gold a number of processes were developed and they can be divided into three groups (Figure 1)."

Jan 1, 2016

By A. Probst, J. A. Finch

"Solids concentration (% solids) profiles and the clear liquid/slurry interface level (mudline) were determined from conductivity measurements in thickeners at Falconbridge' s Strathcona mill (Ni concentrate and non-mag thickeners) and Kidd Creek Division (Cu, Zn concentrates and final tails in the concentrator and thickeners in the Zn plant).The conductivity measurement system consisted of a probe 1.5 m in length with 24 electrode cells, a 24-channel relay, a conductivity meter and a computer. A conductivity profile is collected which locates the mud-line by the sharp change in conductivity. The conductivity measurements are readily converted to solids concentration profiles without need for calibration.The results from the conductivity measurements were compared with those from pressure profiles and direct measurements on local samples obtained using a modified Noranda sampler.It is concluded that the conductivity technique is applicable to locate the mud-line and to estimate the solids concentration profiles in thickeners. This on-line information could improve the operational efficiency of industrial thickeners.INTRODUCTIONAn automated, on-line technique for detecting the clear liquid/slurry interface (or mudline) and the variation in solids concentration with depth would enhance thickener operation and control. In some processes such enhanced control offers significant benefits. For example, in hydrometallurgical operations solid/liquid separation is often the process limiting step. Water treatment plants are increasingly expected to operate remotely and reliable sensing is vital.Current attempts to detect the mud-line include reflection of some signal (e.g. sonar), floats (Dahlstrom, 1985) and measurement of pressure (Rosenblum, 1991), and in one case, natural radioactivity of the solids (Johnston and Simic, 1991). A recent commercial development uses reflection from a vertical array of optical transmitters/sensors, which also gives local solids concentration (BTG Inc., 1991)."

Jan 1, 1993

By Chao Yang, Liguo Tian, Yan Yang

"The Jinfeng gold mine is one of the three gold mines that Eldorado Gold has in China and it is the second largest gold mine in China. The mine is located in a relatively remote and mountainous area in the southwestern Guizhou province. The orebody is similar to the Carlin type deposit in Nevada. The mill throughput was designed to treat 3,600 tones per day. The ore comes from open pit mining and underground mining, and is fed to one-stage jaw crushing. The crushed ore is stockpiled, reclaimed and then fed to the SABC grinding circuit. After classification with hydrocylones, the ore is floated, and the resultant concentrate is oxidized biologically in a process known as BIOX®. The oxidized concentrate is then washed in a CCD circuit and finally leached with cyanide in a CIL circuit for gold extraction. Jinfeng produces about 180,000 oz of gold per year.This paper introduces operations and modifications of SABC circuit over the past three years, and then describes some technical characteristic of the SAG mill and other circuits in the processing plant.Major modifications and improvements on the SABC circuit are described in the paper, including:1. Impacts of grinding ball and mill liner with respect to material selection, quality, hardness, manufacturing methods, and interactions between mill liner and grinding ball.2. Effect of milling operation parameters on milling efficiency.3. Impacts of SAG mill scats and modifications of grate size to improve mill throughput.4. Design and construction of a pebble crushing system, and recycle the crushed scats to the SAG mill to improve mill throughput.With the above-mentioned modifications, 20% increase over the design in mill throughput has been realized over the past three years, and operating cost has been significantly reduced. However, some issues still remain, and thus process modifications and technical innovations are still needed."

Jan 1, 2012

By Garry Biles, Ken Major, Raj Anand

"Prime Resources Group Inc. (Prime) commenced the Eskay Creek Mine operations in 1994 in northern British Columbia. The ore from the underground mine is crushed and shipped to smelters in Japan and Canada. The nominal production rate is 300 tonnes per day.Before commencing the above operations, preliminary evaluations for the high grade gold mine had indicated that an extremely complex metallurgical and chemical process flowsheet would be required to recover the economic metals. Subsequent to the metallurgical testing program, Prime identified markets for the untreated ore and initiated underground mining and ore transportation operations.Prime recently identified lower grade ore zones, which are less economic or uneconomic to mine and ship directly to the smelters. Test programs were initiated to evaluate processes for upgrading the ore to a marketable product. In 1997, Prime will complete the engineering and construction of a 150 tonne per day flotation plant to concentrate the gold and silver values for the NEX and 109 zones. The mine will then, in addition to the direct shipping of the shippable ore, ship concentrate from the millable ore. This paper discusses the development and design of the flotation plant.IntroductionThe Eskay Creek Mine is located in north-western British Columbia approximately 83 km north-west of Stewart, B.C. as shown in the location map Figure 1. Access to the site is provided by a restricted access gravel road from Bob Quinn Lake junction of highway 37, a distance of about 59 km. The work force at the mine is based on a two-week rotation schedule. Employees are flown from as far away as Vancouver to Smithers by commercial airlines and then from Smithers to the Bob Quinn airstrip by local charter. During poor weather conditions, transportation to site is by vans that take about six hours to travel the 450 km road distance."

Jan 1, 1998

By André Desbiens, Frédéric Flament, Lionel Ryan

"In December 1991, the GRAHM (Groupe de Recherche sur Jes Applications de L'informatique a L'industrie Minerale), of Laval University started a KBAC (Knowledge Based Automatic Control) grinding project with Falconbridge Ltd. The project was shifted to the Kidd Creek concentrator in the summer of 1992, with the first phase involving the collection of data by GRAHM and Falconbridge personnel for the construction and tuning of a dynamic simulator. This grinding circuit dynamic simulator known as DYNAFRAG was used to assess various control strategies, with a PID-based strategy being selected for ease of installation on a Foxboro II A DCS.The ""B"" primary grinding circuit was chosen for the initial strategy development, which is one of three parallel divisions. The KBAC project was successfully completed in November 1994, with subsequent modifications being implemented by Kidd Creek staff.This paper summarizes, the current control strategy, the project costs, and the realized benefits. Some of the problems which were encountered with the initial installation, and the subsequent modifications are also dealt with in this paper.2.0 IntroductionIn the summer of 1992, the GRAHM of Laval University and Falconbridge Ltd. started a project to design a control strategy for the ""B"" grinding circuit at the Kidd Creek Division. Between 1992 and 1994 several papers have been written to describe the control objectives, the methodology followed, and some initial results (Flament, 1994A, 1994B). Since then, however, the control objectives changed, and several improvements were made to the circuit. In particular, maximizing tonnage was no longer a priority, a variable speed drive pump was installed on the cyclone feed, the ball mill power was lowered, and the geometry of the cyclones changed. As a consequence, the control strategy was revised. Table 1 summarizes the historical developments on the ""B"" division grinding control."

Jan 1, 1996

By K. Mayhew, A. Lossin, H. Salomon-de-Friedberg

"The proportion of arsenic-bearing copper concentrates that supply the copper industry is on the increase. The current strategy of miners and smelters to blend feeds to reduce average arsenic levels is rapidly becoming unsustainable as more concentrates with even higher arsenic levels enter the marketplace. In addition, the capability of smelters to manage increasing arsenic levels to more stringent and restrictive environmental standards is limited by available process technologies.CESL Cu-As Technology is a viable technical and economic alternative to traditional smelting. It offers the advantage of simultaneous leaching and precipitation of the arsenic in a single processing step while achieving high copper extractions. The process has been piloted in six separate, fully-integrated pilot campaigns for feeds ranging from ~1% to 12% arsenic in concentration. This paper will focus on the behaviour of arsenic in CESL’s leaching/precipitation step and operational factors that influence it, and on the chemical and physical characteristics of the subsequent iron-arsenic residue that represents today’s best available technology from a long-term sustainable (capital/operating costs, metal recovery, safety, environmental) perspective.The examples provided will demonstrate copper extractions approaching 99% and arsenic deportment to residue of 99%. Instantaneous settling rates of the residue were measured at 15+ m/hr. Data from subsequent short and long-term stability testing of the residue confidently support the view that the scorodite-bearing residues from the process will be stable and provide an acceptable arsenic management practice. The technology offers increased revenue and environmental certainty for miners, smelters & refiners to support the development of metallurgically challenged copper-arsenic deposits. Furthermore, results will be presented demonstrating the technology’s ability to successfully incorporate hazardous arsenic containing smelter by-products such as dust into the process feed and simultaneously extract copper value while fixing the arsenic in a stable residue."

Jan 1, 2016

By Robert W. Baker

"Brunswick Mining and Smelting commenced production at the No. 12orebody, a lead/zinc/copper/silver deposit, situated some seventeen miles southwest of Bathurst, New Brunswick, in 1964 at 4100 tonnes per day. Today it operates at 10500 tonnes per day.Brunswick has utilized underground incentives since commencement of operations. Early incentive systems were “off the wall"" based on those in effect at other properties then manipulated to fit, often hack calculated from actual performances to tm1ve at the desired result. If subsequent payouts were considered too high the rates were simply reduced, often retroactively. A large amount of this was&apos; subjective and depended to a large extent on whoever was handling the bonus, in conjunction with senior mine department staff.In 1971, it was&apos; decided that a more scientific approach was needed, particularly with the introduction of the new cut-and-fill mining. After some research, a system based on Work Study was selected. Another major mining company employing such a system loaned one of their experienced employees to train some Brunswick personnel in the art of Work Study and conversion of data to incentive rates, and in 1972, the first such developed incentive rate was-introduced.Along the way, borrowing from others where necessary, a set of Brunswick principles and parameters were developed. While refined to some extent, these remain basically the same today.Improvements in methods and equipment have resulted in improved performance and reduced costs. The use of underground incentives has played an important role in this. A/though individual bonus earnings have steadily increased, bonus cost per tonne has remained nearly constant - in real dollars, it has regularly decreased. These are two of the main parameters of a successful system."

Jan 1, 1995

By T. E. Newkirk

In the past few years there has been considerable controversy on the most suitable material to be used in ventilation doors which will have to serve also as fire doors. For this reason it was thought useful at Mclntyre to find out to what extent our zinc chloride treated spruce doors were fire resistant and what protection we could expect in case of an underground fire. The doors to be tested were constructed of the same material as that used in our regular ventilation doors underground, namely, 2-inch tongue-and-groove well seasoned spuce. It is to be noted, however, that whereas the dimensions of the ventilation doors are 5 ft. by 6 ft., those of the test doors were only 14 in. by 17½ in. The treated doors were made up from lumber that l1ad been treated with chromated zinc chloride, as described in Mr. Keeley&apos;s paper. Two kinds of tests were carried out: a heat test and a flame test. An electric muffle furnace for assaying, equipped with a pyrometer, was used for the beat test, the furnace door being replaced by the wooden door to be tested. In the flame test, the door to be tested formed one end of a rectangular steel box in which was placed a large blow-torch. Air was admitted at the rear of the box. A pyrometer recorded the temperature at the inside surface of the door where flame impinged. The results of the tests are presented in the accompanying table, in which the word &apos;failed&apos; is to be interpreted as meaning that the door had been damaged to the extent that smoke, but not necessarily flame, issued from its outside surface.

Jan 1, 1949

By E. L. Jones

THE concentration of tin in the tailing from Sullivan ore has been de-scribed by H. R. Banks, superintendent of the Sullivan concentrator, in a paper presented at the annual meeting of the B.C. Division of the Institute in October, 1941(1). The reduction of the concentrate EARLY EXPERIMENTAL WORK Experimental work on the smelting of Sullivan tin concentrate was done in Trail. Preliminary tests in the laboratory indicated that the concentrate could be successfully treated in an electric furnace. A small electro-thermic pilot plant was built and, at the same time, reduction tests were made in an oil-fired reverberatory furnace. Both types of treatment gave satisfactory results. It developed that no fluxes, other than those already in the ore and in the ash of the reducing agent (either coke or coal), were needed. By using an excess of iron, a complete recovery of the tin in the slag was obtained, but with 20 to 30 per cent iron in the tin. A counter-current treatment was thus indicated, viz., the primary slag was first treated with hard-head (iron-tin alloy) from the refining and final slag smelting, producing tin with 2 to 4 per cent Fe. This secondary slag was subsequently treated with scrap iron to produce tin-iron alloy, and a final slag low enough in tin to waste. By following this line of endeavour, a tentative flow-sheet was drawn up and a successful operation was established. Without resorting to electrolytic refining, a metal comparable in purity to standard tin was produced. Metallurgical results and operating costs were found to be very similar with an electric furnace or a fuel-fired furnace. &apos;However, electric furnace smelting was chosen because the furnace required was smaller, and because some doubt existed as to the cleaning of the final slag in the reverberatory furnace reduction.

Jan 1, 1944