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By J. D. Little

"THE SALMO PROPERTIES of Canadian Exploration, Limited comprise forty-one Crown-granted mineral claims situated on Iron mountain between Sheep creek (on the north) and Lost creek, some eight miles southward from Salmo, in the Nelson mining division (see Figure 1). On these claims there are at present four operating mines. These are, from north to south, the Dodger, Feeney, and Emerald, all tungsten mines, and the Jersey, a lead-zinc mine, about 7,000 feet south of the Dodger. About midway between the Dodger and the Jersey is a fifth mine, the Emerald lead-zinc mine, which was the first producer in the area but has been inactive since 1926. The locations of the mines are shown in Figure 2.Records of prospecting and claim staking on Iron mountain -which owes its name to the presence of large iron-stained outcrops -go back to 1895. The first recorded owner of the claims that form the old Emerald lead-zinc mine was John Waldbeser who, in 1908, did considerable work and produced 426 tons of lead ore valued, at the then price of lead, at $7,000. This ore was chiefly lead carbonate and was taken from the surface outcrops of two bands of ore that form the Emerald orebodies. In 1909, the Emerald was the largest lead producer in the Nelson mining district."

Jan 1, 1953

By Dennis Z. F. Mraz

"Because rock salt and pot ash formations contain particles capable of elastic deformation, the material has usually been compared with various rheological models containing elastic elements. Extensive laboratory testing of salt and potash samples resulted in various theoretical conclusions based on conditions quite different from those in a deep mine. It is felt that sufficient data have now been collected from underground measurements and observations to demonstrate the inelastic behaviour of salt masses in a deep mine. The conditions in deep salt strata can be defined as near hydrostatic, because the acting pressures are very high compared to the negligible elastic limit of the salt. The pressure field and creep around the single opening is a function of its shape, size and age. The same factors, together with extraction, determine the behaviour of the mining a rea. The application of the correct analytical method in mine design can significantly improve the excavation techniques and predetermine the long-term behaviour of openings."

Jan 1, 1973

By J. Font

The RAF or copper fire refining process at the Caletones smelter of CodelcoChile consists of several pyrorefining stages that successively lowers the levels of S, Fe, Pb, As, Sb, Te, Se and to some extent 0 in the copper melt, allowing the production of high quality fire-refined copper of 99.93 % purity (RAF copper). This chemical feature of the RAF copper is only achieved by quite long processing times, and a considerable consumption of fluxes and fuels in the RAF vessels. The main goal of this study was to improve the RAF process performance by thermodynamic considerations and industrial scale tests, i.e., an in situ approach for the process as featured by intensive sampling, characterization and subsequent modifications. Thus, both theoretical and industrial data were derived, and based on kinetic and thermodynamic evaluations, a higher operational standard was achieved for the RAF process. These new operational practices allowed enhancing the overall economic of the RAF process, by reducing the time and operational costs, and consequently increasing the availability of the vessels.

Jan 1, 2007

By W. A. Jankola, M. T. Martin

"The world's first zinc pressure leaching plant at Cominco's Trail operations started production in 1981. Sullivan zinc concentrate is leached at the rate of 190 tpd in a four-compartment autoclave at elevated temperatures and oxygen pressures to produce a zinc sulphate solution and elemental sulphur. Zinc extraction exceeds 98% and sulphide to elemental sulphur conversion exceeds 95%.The sulphur is recovered from the discharged slurry in the underflow of a hydrocyclone, and is upgraded as a flotation concentrate. The concentrate is dewatered, melted and filtered to produce a polished sulphur product of better than 99% purity. An alternative sulphur recovery process employs an unique sulphur separation method that recovers over 90% of the sulphur in the molten state. The remaining sulphur is recovered by flotation following flash discharge of the slurry. The product slurry containing the zinc sulphate solution and the iron-lead residue joins the existing roast-leach zinc plant for purification and zinc electrowinning.IntroductionIn 1981, Cominco Ltd. commissioned the world's first commercial zinc pressure leaching plant at their Trail operations in British Columbia. As part of a $700 million modernization and expansion program for lead and zinc started in 1977, the company installed the first o f two autoclave trains.Prior to pressure leaching, zinc production was closely tied to acid and fertilizer production. Sulphur dioxide produced from roasting the concentrates had to be converted to sulphuric acid. Thus, any expansion of zinc operations with the roast-leach technology would have required extensive modifications to the acid and fertilizer plants as well as to the roasters. Zinc pressure leaching, on the other hand, offered a one-step dissolution of zinc from the concentrates with high extraction, and production of elemental sulphur. The associated dust and sulphur dioxide generated in the roasting process are both eliminated from the working environment of the pressure leach facility. The flexibility allowed by producing elemental sulphur, and the opportunity to break the sulphuric acid link for about 25 per cent of the targeted 272,000 tonnes per year zinc production was very appealing. The process was successfully demonstrated at a pilot plant in 1977(1-6)."

Jan 1, 1985

By The Hollinger Mill Staff

MR. E. L. Longmore (l), in presenting this paper at the Annual General Meeting, suggested that it was neither desirable nor necessary to read any but the briefest portion of the paper. He wished to emphasize that the scale of the milling operations at the Hollinger not only provided several units in continuous operation to ensure that the test work was being carried on under normal conditions, but also provided a large and competent staff to supervise such an extensive experiment and to criticize intelligently the recorded results. He wished to acknowledge the enthusiastic co-operation of the group of assistants whose excellent work was responsible for the confidence with which the conclusions could be presented. The following data were given as a supplement to the published paper. Following conclusion of the test work described in the paper, the size of the balls fed to the mill was changed from 3-in. diameter to 2Yz-in. Within a short time, a gradual rise in the tonnage rate took place and, as no other condition had been changed, the increased tonnage was credited to the changed character of the ball load. At present, the average daily tonnage for the mill with low discharge is 1,507 tons, an increase of 77 tons over the 1,430 tons ground when feeding a 3-in. ball. A similar increase in tonnage should be credited to the high-discharge test, if the 2;Y2-in. ball had been used at that time. This would raise the 1,117 tons ground at that time to a new figure of 1,177 tons. From a study made to determine those factors influencing the rate of pulp flow through the low-discharge mill, it became evident that the ball screen used on the high-discharge mill must have been restricting the flow through the mill, although this was not recognized at the time. Accord-ingly, it was decided to repeat the high-discharge run. Additional length was added to the 21-in. ring grizzly originally used, and the length of the reverse spiral was increased. The end of the grizzly was left open so that, if necessary, excess pulp might flow out. A coarse mesh 3 ft. by 3 ft. flat screen was provided to catch stray balls.

Jan 1, 1937

By M. Drozd, M. Somppi, Z. F. Yamak

"Activated carbon is used to recover gold in solution from various gold circuit designs including carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) processes; however, the characteristics and durability of the activated carbon can vary. Handling this medium poses challenges when fine carbon losses occur in the CIP, CIC, and CIL circuit discharge. Additionally, adsorbed gold on carbon fines can be lost to the discharge. The studies in this paper aim to standardize laboratory methods for testing activated carbon qualities specific to the gold industry. The investigation focused on size distribution, platelet content, wet attrition resistance, gold kinetics, and gold loading. The results demonstrate that a 96 h attrition test and a gold loading test form a basis for ranking carbon samples based on expected economic performance of the plant. ¦ KEYWORDS Activated carbon, Carbon attrition, Carbon ball-pan hardness, Carbon evaluation, Carbon fines loss, Carbon turnover RÉSUMÉ Le charbon actif sert à récupérer l’or en solution dans diverses conceptions de circuits de broyage d’or, notamment les procédés de charbon en pulpe (CIP, de l’anglais carbon-in-pulp), de charbon en colonnes (CIC, de l’anglais carbon-in-column) et de lixiviation au carbone, ou charbon au lixiviat (CIL, de l’anglais carbon-in-leach). Toutefois, les caractéristiques et la durabilité du charbon actif varient. La manipulation de ce milieu pose des problèmes dès lors que se produisent des pertes de carbone à grain fin dans la décharge des circuits CIP, CIC et CIL. Par ailleurs, l’or adsorbé sur les fines de charbon peut se perdre durant la décharge. Les études réalisées dans le cadre de cet article visent à normaliser les méthodes de laboratoire pour tester les qualités du charbon actif spécifiques au secteur de l’exploitation aurifère. L’étude portait sur la composition granulométrique, la teneur des plaquettes, la résistance au frottement humide, la cinétique de l’or et l’adsorption de l’or, à savoir la capacité du charbon actif à adsorber l’or sur une période spécifique. Les résultats ont montré qu’un test de frottement de 96 heures et un test d’adsorption de l’or constituaient le fondement d’un classement des échantillons de carbone du point de vue économique en vue de trouver l’échantillon générant le meilleur rendement des usines. ¦ MOTS CLÉS charbon actif, dureté de la batée des billes de carbone, évaluation du carbone, perte de fines de charbon, renouvellement du carbone, résistance au carbone INTRODUCTION The use of activated carbon for gold extraction is among the most common processes in modern gold-milling plants. Although gold losses are inevitable in the milling process, operators often focus on reducing gold loss to tailings. One focus area that is important in gold recovery is optimizing the use of activated carbon and examining potential changes to realize improved economics for carbon-in-pulp (CIP), carbon-in-column (CIC), and carbon-in-leach (CIL) processes. Granular activated carbon (Figure 1) is widely used for gold extraction from slurries containing gold-cyanide complexes. This activated carbon—produced from charred coconut shells—is well suited because of its adsorption kinetics, gold-loading capacity, attrition resistance, and ability to be reactivated. The extraction process using activated carbon technology involves adsorbing gold from an aqueous leachate; the gold is dissolved using a cyanide salt and consequently a dicyanoaurate(I) ion (Au(CN)2–) is formed, as noted in equation 1."

Jan 1, 2019

By G. H. Montgomery

THE PROPERTY of Aunor Gold Mines, Limited, consists of eleven claims, located in Deloro township, approximately three miles southeast of the town of Timmins. Production began in 1940 and currently the mine is operating at a rate of 485 tons per day. GEOLOGY AND MINE DEVELOPMENT The ore zone at Aunor occurs in altered lavas bounded on either side by soft talc-chlorite schist, locally known as 'soapstone'. The zone has a strike of N.80°E, dips 60° North, and varies in width from 0 to 250 feet. Within this zone the main orebody has a westerly plunge, away from the No. 1 shaft area (Figure 1). The mine is developed by a vertical three-compartment shaft from surface to the 3,082-foot horizon. Working levels are spaced at 125- foot intervals from the 625-foot to the 1,625-foot horizon, from which to tl1e 2,125-foot level the interval is 250 feet.

Jan 1, 1957

By N. R. Andrade-Gonzalez

The present work was performed on four alloys containing Al-6 2 wt pct Si-3 5 wt pet Cu, with magnesium in the range of 0.02-0 40 wt pct, and strontium in the range of 0-165 p p m Tensile bars were cast in a permanent mold (ASTM B108-160) and heat treated at 495°C for 4 hours, followed by quenching in hot water. Two different artificial aging temperatures were applied at 180 °C and 220 °C for times from 0 5 h up to 8 h Tests were conducted with the intent of determining how the mechanical properties of an aluminum-based 319 alloy are affected when aging time and temperatures are varied In general, it is seen that the alloy does not exhibit the common peak-overaged aging condition due to the presence of several coexisting phases, indicating that they form through complex eutectic reactions late in solidification.

Jan 1, 2004

By J. R. Giegerich

THE problem of dust control in metal mines has, of late years, received considerable attention. In 1932, a detailed study of dust conditions at the Sullivan mine was undertaken and, while there are no actual data available regarding conditions prior to that date, there has unquestionably been a great improvement as a result of the investigation and of the preventive methods adopted. In attacking the problem, three main factors were considered: first, and most important, how to prevent dust from being formed; second, how to depress or collect it if formed; and third-if unable to accomplish these-what precautionary measures must be adopted to prevent inhalation of this dust. Seven distinct methods of dust control are now practised at the Sullivan. Six of these are employed underground; they are tabulated below in the order of their relative importance. The seventh method is used in the crushing plants on the surface. (1) Improved ventilation (2) General use of water sprays and atomizers (3) Curtailment of secondary blasting during shift hours ( 4) 100 percent wet drilling-primary and secondary (5) Wearing of protective equipment (6) Change from carbide to electric lamps (7) Installation of dust collectors

Jan 1, 1936

By A. L. Valdivieso

"The correlation obtained by plotting pH of maximum adsorption of dextrin versus the isoelectric point of involved metal hydroxides can be used to predict the depressing effect of dextrin on flotation in ""pure"" sulfide mineral systems. Sulfides are strongly depressed by dextrin whenever their surfaces are coated by the corresponding metal hydroxide. The efficiency of dextrin as a depressant is affected by the adsorption of the sulfhydryl collector. Hydrolyzable cations modify this relationship in flotation of sulfide ores. This type of deviation of the real-life systems from the predictions based on the behavior of the pure systems shows the possible ways that could be followed to further improve the obtained correlations. Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists ofCIM Sudbury, Ontario, Canada Edited by"

Jan 1, 2009

By J. A. Finch

The metallurgical performance of a flotation cell is strongly influenced by the features of the bubble population, characterized by the gas dispersion properties, a collective term including bubble size, superficial gas velocity and gas holdup. It is required that gas be well distributed throughout the cell. Gas distribution is estimated here by simultaneous measurement of gas velocity (Jg) at different distances from the center (radial Jg mapping). This communication describes a technique to perform radial Jg mapping together with fundamentals for the interpretation of the data.

Jan 1, 2009

By C. C. Walden, D. W. Duncan, P. C. Trussell

Thiobacillus ferrooxidans released copper from -400-mesh chalcopyrite at a rate of 54 mg/I/hr. Any flotation chemicals remaining on the concentrate did not inhibit the leach rate. Zinc rougher tailings were also leachable. With bacteria, 75 per cent of the zinc and 68 per cent of the copper came into solution. In the sterile controls, only 16 per cent of the zinc and 10 per cent of the copper were released; 100 per cent of the copper and 36 per cent of the zinc were leached from pyrite cinders through the action of bacteria, and 17 per cent of the copper and 10 per cent of the zinc were solubilized in the sterile controls. Less than 0.027 per cent of ~he iron came into solution. The majority of the copper in a converter slag was acid-soluble, but a reverberatory slag responded to microbiological leaching. About 62 per cent of the copper came into solution, compared to 12 per cent in the sterile control. The various factors influencing the rate and extent of microbiological leaching are discussed.

Jan 1, 1966

By S. D. F. Rocha1, L. Katsarou, G. P. Demopoulos

"Crystalline scorodite (FeAsO4·2H2O) has been pointed out as the most stable compound to immobilize arsenic for disposal due to its good stability performance according to EPA’s TCLP test. However, in anaerobic conditions, arsenic releasing is increased to undesired levels due to scorodite reductive break down, in eh <100 mV at 5.5< pH <7 which causes contamination of groundwater. In this work, the concept of encapsulation by controlled deposition of mineral coatings immune to pH or redox potential variations is presented. Aluminum and calcium phosphates were evaluated as coatings and the release of arsenic from scorodite previously coated with phosphates was substantially decreased when compared to the VALUES (obtained from crude scorodite, both in oxic and anoxic conditions. INTRODUCTIONThe widely acknowledged hazardous properties of arsenic as a soil and natural water contaminant and its adverse health effects when present in drinking water have led to very strict environmental regulations (Mandal & Suzuki, 2002) aiming at the drastic reduction of the presence of arsenic in the effluents of the mining and metallurgical industries.In Canada, the Metal Mining Effluent Regulations (Government of Canada, 2002) dictate that the maximum monthly mean concentration in tailings for arsenic must be lower than 0.5 mg/L. At the same time the new Environmental Protection Agency’s (EPA) standard for drinking water foresees concentrations of arsenic less than 0.01 mg/L, which means that the leachability limit for the industrial waste is set at 1 mg/L arsenic (EPA, 1994; Le Berre et. al, 2007). One of the preoccupations of the mining and metallurgical industries is directed to the elimination of arsenic presence in all its effluents, especially since arsenic is commonly found in many ores mainly in the form of oxides and sulphides. After processing of arsenic-containing ores and concentrates, arsenic can be present either in solid residues and tailings, liquid effluents or in flue dusts. Several of these forms of arsenic waste materials are soluble and therefore may cause the contamination of ground water, if disposed of without further treatment."

Jan 1, 2012

By A. S. H. Kabir

Due to the high specific strength and excellent corrosion resistance, Ti-6A1-4V has been widely applied in aerospace industries. In this study, the welding performance of 3.2-mm and 5.1-mm thick Ti¬6A1-4V sheets was studied using a 4 kW continuous wave Nd: YAG laser. It is found that the use of filler wire, matching the parent metal composition, can bridge the joint gap and produce full penetrated welds up to a width of 0.6 mm without cracking. The laser welds were characterized in terms of the bead geometry, defects, microstructures, and hardness. With increasing joint gap, the percent porosity area increased, reaching just over 1% of the fusion zone area at a gap width of 0.6 mm for the 3.2 mm and 5.1 mm thick sheets. The maximum underfill depth in the Ti-6A1-4V laser welds was about 5% or 7% of the sheet thickness for the 3.2-mm and 5.2-mm thick materials, respectively, meeting the AWS D17-1 specification. The microindentation hardness was maximum in the fusion zone and sharply decreased through the heat affected zone until reaching the base metal value.

Jan 1, 2011

By A. W. Jacob

THE shatter cut, or burn cut as it is often called, is one in which the centre of the round is almost reamed out by drilling holes so close together that the blasting of a few holes shatters the rock embraced by all of them. The idea of shattering the centre of a face by shooting holes drilled parallel to one another is not new. Some miners have used this method for years past, when drilling quartz or a face near timber. But, they followed the age-old practice of first blasting the cut, then the square-up. It was not thought possible to shoot both together. In the case of machine men, the two largest contributors to silicosis are dust from drilling and dust from blasting. It follows, then, that material reduction of exposure to either constitutes an important step in lessening the incidence o silicosis. It was? the hope of being able to do away with the necessity of the men having to return early to a face, through dust concentrations of 60,000 to 80,000 particles per cubic centimeter, that aroused interest "in the shatter cut at McIntyre. Efforts to this end were successful. Now, with this kind of cut, all holes of a round are shot simultaneously.

Jan 1, 1942

By Gillyeard J. Leathley

The Cassiar Asbestos mine is situated on McDame Mountain in northern B.C. at a mean elevation of 6000 ft (1830 m), some 6 miles (9.6 km) from the townsite /mill complex. Asbestos ore production averages 1,200,000 tons at 9% R.M.G. (Recoverable Mine Grade) and waste production is 6,000,000 cu. yd annually. Waste is loaded by P&H 1900 AL 11-cu.-yd shovels and transported by 75-ton and 85-ton end-dump trucks to the waste dumps. The ore is loaded by a 992 Caterpillar 10-cu-yd front-end loader into 50-ton and 75-ton trucks, which transport it to the crushing plant located at the mine. The ore is transported by a 15,370-ft (4685-m) tramline to the mill complex.

Jan 1, 1978

By E. Rosseland

Glencore Nikkelverk is operating its Chlorine Leach Process in Kristiansand, Norway, producing nickel, copper and cobalt metals by electrowinning. The copper tankhouse, with an annual capacity of 40 000 tons, is based on traditional technology with copper starting sheets and cast lead anodes. Due to the integration of this copper sulphate process in a nickel chloride refinery, operating conditions are unique. In recent years, operating costs have escalated, and health, safety and environment are top priorities. Test work on a new copper process involving permanent cathodes was therefore initiated in 2007, and a pilot plant was started up in 2012. The objective was to introduce innovative technologies and tailor these to be robust in the Nikkelverk process, to form a basis for a new sustainable copper tankhouse with a high degree of automation. Key learnings from more than 10 years of R&D work is presented, focusing on cathode materials and copper metal deposition including product quality, robotic stripping, coated titanium anodes, electrode contacts and alignment, electrode current monitoring with Hatch HELM tracker technology and acid mist capture using cell hoods with off-gas scrubbing. The new copper tankhouse project was approved in December 2018, and commissioning is planned in 2022. INTRODUCTION The Kristiansand nickel refinery was started up in 1910, processing nickel-copper matte originating from Norwegian mines. It was the first industrial plant in the world to produce nickel by electrolysis using the Hybinette electrorefining process. After roasting to burn off sulphur in the matte, copper was leached and recovered as cathodes by electrowinning with lead anodes. In the first years, the copper cathodes were fire- refined and cast into ingots for sale (Thonstad Sandvik, 2004). The Nikkelverk refinery was acquired by Falconbridge Ltd. in 1929, and since then matte from Sudbury, Canada has been the main feed source. The Hybinette process was in operation until 1975 when conversion to the newly developed Chlorine Leach Process was started (Stensholt, Zachariasen, & Lund, 1986). In this process still used today, nickel and part of the copper in the matte is leached using chlorine gas generated on dimensionally stable anodes in the nickel and cobalt electrowinning tankhouses, as illustrated in Figure 1. The presence of dissolved copper ions in the strong chloride leach solution is essential for effective chlorine absorption and nickel dissolution. To improve copper-nickel separation, slurry from the atmospheric leach is pumped to autoclaves and then to cementation tanks for precipitation of copper as copper sulphide, which is collected in filter presses. The leach residue, also containing the precious metals, goes through a 3-stage wash to remove chlorides before being introduced as a slurry to fluid-bed roasters for sulphur removal. The resulting calcine is then leached in spent electrolyte from the copper electrowinning tankhouse, to dissolve most of the copper and also some nickel, cobalt and other elements. After separation of the leach residue, elements like Bi, Te, As and Ag are removed from the strong copper sulphate solution by reaction with metallic copper in three columns in series filled with cut copper cathodes. The purified solution then passes polishing filters before being mixed with spent electrolyte and returned to the electrowinning cells.

Jan 1, 2019

By J. Jacob, A. G. Guézennec, F. Savreux, S. Nanaa, M. Delclaud, P. d’Hugues

Mining wastes can contain base and precious metals, but also metalloids and rare earth elements that are nowadays considered as highly critical for the industrial development of the European Union. The development of alternative routes to conventional processing is still required in order to decrease the cost associated with the treatment of these resources, which are more complex in composition and with lower grades. An ecologically acceptable and yet economic alternative for processing of low-grade sulphidic primary ores as well as mining wastes is the extraction of metals using bioleaching processes. The purpose of this study was to test the potential of bioleaching procedures for the extraction of residual metals contained in sulphide-rich tailing wastes coming from a European copper mine. These wastes are mainly composed of pyrite (60%) and contain cobalt (0.06 %) and gold (0.95 g/t). Bioleaching experiments were performed in 2 L stirred reactors in batch mode and in pilot tanks (110L) in continuous mode, using the "BRGM-KCC" bacterial consortia. High metal extractions were obtained (cobalt dissolution > 90%), which confirm that bioleaching is an effective process for this tailings material. Experimental results were used to design a global treatment process scheme and to perform a preliminary economic evaluation, showing good profitability of a hypothetical industrial project.

Jan 1, 2014

By D. Q. Li

"This paper briefly introduces problems that can occur in the design of rare earth separation processes and that require close attention during engineering design. Important factors must be considered include the following. 1) The separation process is primarily based on two points. One is the raw material composition and the other is the product purity requirements. 2) Optimization of the rare earth separation process requires optimization of the flowsheet configuration and each process parameter. 3) The separation of heavy rare earths is performed under non-equilibrium conditions. 4) The radioactive elements uranium and thorium must be removed from the feed to separation. 5) Process economy and environmental protection must be considered together in the evaluation of any separation technology options. INTRODUCTIONThe importance of rare earth (RE) is well recognized in high-tech applications and sophisticated national defense technologies. RE is the treasure house of the 21st Century new materials. The RE separation processes developed using HEHEHP (P507), P507-alcohol (ROH) and naphthenic acid (NA) have been widely applied in China. These processes have led to industrial production of all of the individual RE with up to 4N (99.99%) or higher purity (Li, Wan, Lin, Xie, Lin, & Wang, 1980; Li, Wang, & Xie, 1985; Li, 1995; Li, Wang, Meng, & Li, 2005). In order to achieve the best economic and environmental benefits, this paper discusses several problems to which attention must be paid in the design of rare earth separation process."

Jan 1, 2016

This paper describes results of an investigation of fracture during tube expansion of 5052 and 5754 aluminum seam-welded tubes to optimize the forming of aluminum tubes during hydroforming. The tubes were produced in laboratory by press forming of tubular blanks and pilot-scale high frequency induction welding. The effect of property mismatch between the weld and base metal has been investigated. The material work-hardening was obtained by roll forming of sheet blanks in the 0-temper and H32 temper condition. The results showed that a significant work-hardening can be present in aluminum tube without causing weld fracture during tube expansion. The work-hardening increases the yield-strength mismatch between base metal and weld. Annealing reduces the hardness of base metal leading to more uniform deformation during tube expansion. As a result the occurrence of weld fracture is reduced. The study showed significant deformation of the welds during tube expansion. Weld thinning can be as large as parent-material thinning before fracture. To avoid fracture at the weld joint, good formability and high yield strength of the weld is needed. FEM simulation showed that minimizing the width of the weld joint can reduce thinning of the weld.

Jan 1, 2004