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By J. G. Benford, E. B. Stanley

THERE have been numerous papers in the literature1 on the subjects of secondary and tertiary recrystalli-zation in silicon steels. Secondary re crystallization is believed to occur in silicon steel by either a surface-energy phenomenon such as in the usual development of cube texture or by a second-phase phenomenon such as in the usual production of cube-on-edge texture, (110)[001]. Tertiary recrystallization in silicon steel is thought to be a surface-energy phenomenon in which, for example, grains having a (110) plane parallel to the sheet grow at the expense of other grains in the matrix including those having a (100) plane parallel to the plane of the sheet. In all these instances, the orientations involve either a preferred (100) or (110) plane in the plane of the sheet. Until very recently, there had been no experimental evidence reported of the occurrence of a texture by secondary or tertiary recrystallization that involves a (111) plane in the plane of the sheet, although Kohler reported that grains having the (111) plane parallel to the sheet surface can also grow by a surface energy mechanism. Stickels observed a secondary texture in iron described as "near {554}(225)" about 5 deg from {lll}(211).3 Mee reported that in a vacuum-remelted commercial 3 pct Si steel {lll}(hkl) secondary recrystallization took place in a high vacuum and in pure hydrogen.4 Recently in our Laboratory, large-grained structures such as shown in Fig. 1 having an orientation near (111)[211] or (111)[2ii] have been observed. The arrows point to areas of relatively small matrix grains of the kind that the much larger (111) grains consumed while growing. It is of interest that the matrix grains of samples annealed at 2000°F were essentially the same size as those in samples annealed at 2200°F, a matrix-grain-size stability that is characteristic of secondary recrystallization processes. A torque disc taken from Grain 4 of the sample shown in Fig. 1 revealed a very flat magnetic-torque curve, as expected, with a maximum torque of 7000 dyne-cm per cu cm compared to the typical 180,000 dyne-cm per cu cm observed for conventional (110)[001] textured material. A back-reflection Laue pattern for Grain 4 is shown in Fig. 2 and indicates a near-perfect (111)[211] orientation. The other six numbered grains in the sample of Fig. 1 had a similar orientation. The small circles next to each number were the target areas for the X-ray beam. The starting material was an 0.011-in.-thick oriented silicon steel, (110)[001] texture, containing 3 pct Si. It was lightly pickled, cold-rolled to 0.005 in., and annealed 4 hr at 2200°F within a nickel box (not gas-tight) inside a tube furnace through which dry hydrogen was flowing. The samples were heated to 1500°F in 1 hr and then at 60°F per hr to 2200°F. There were other

Jan 1, 1969

By K. A. Jackson

An analysis of the conditiom fm the dendrite-eutec-tic transition is presented which takes into account a temperature gradient. The composition and temperature in the liquid ahead of a eutectic interface growing at a particular velocity are used to determine whether dendrites can grow faster than the eutectic interface in the liquid which exists ahead of it. If the dendrites grow faster, then a mixed structure with dendrites growing ahead of the eutectic interface is expected. If the dendrites do not grow faster, then only eutectic will be present. The analysis is presented for Sn-Pb system, for which sufficient experimental data to make the analysis are available. The transition from eutectic to dendrites-plus-eutectic derived by this analysis is in good agreement with the recent experimental observations of Mollard and Flemings on Sn-Pb alloys. In alloys which grow with a lamellar eutectic structure, the lamellar structure can be grown over some range of compositions close to the eutectic composition. When the composition is too far from the eutectic composition, dendrites of one or another of the primary phases grow ahead of the eutectic interface, producing a microstructure consisting of "primary phase" interspersed with a lamellar eutectic. Hunt and Jackson' have recently predicted the range of compositions over which the eutectic phase should grow in Sn-Pb alloys, and the composition-growth rate dependence of the transition from eutectic to dendrites-plus-eutectic. This prediction was based on work on organic eutectics. The model on which the prediction was based assumed that there was no temperature gradient in the system, a condition approximately satisfied in the study of organic eutectics2-5 but not in the growth of metallic alloys. Mollard and Flemings6,7 have recently reported experimental results for Sn-Pb alloys in which they have located the transition from eutectic to eutectic-plus-dendritic growth. The experiments were performed using large temperature gradients (65 to 515 deg per cm), so that the assumption of zero gradient on which the predictions of Hunt and Jackson' were based does not apply. It is not surprising therefore that the predictions' were not born out by the experiments.7 In the present work, a modification of the model used by Hunt and Jackson is presented which takes into account the temperature gradient. The modified form of the theory predicts the transition from eutectic to eutectic-plus-dendritic growth in terms of the composition, growth rate, and temperature gradient. Good agreement with the experimental data of Mollard and Flemings7 is obtained. THE HUNT-JACKSON MODEL The model used by Hunt and Jackson' to predict the transition from eutectic to eutectic-plus-dendritic growth was based on the assumption that dendrites grow ahead of a eutectic interface only if they grow faster than the eutectic interface. Analytic expressions for the growth rate of dendrites and the eutectic in Sn-Pb alloys were obtained by fitting experimental data with equations whose form was indicated by theory.

Jan 1, 1969

By James F. Watson, Jack L. Christian

The austenite to martensite solid state phase transformation in the 300 series austenitic stainless steels was studied over the temperature range of 788° to -423°F. The primary variables controlling this transformation are temperature, plastic strain, and chernical composition. Consequently, temperature effects were studied by measurements at 78", -100°, -320°, ad -423°F; plastic strain was controlled by cold-rolling the alloys varying amounts between 40 and 80 pet; and chemical compositions were varied by using 301, 302, 304 ELC, and 310 stainless steel alloys which exhibit varying degrees of austenite stability. The martensite contents of various alloys were correlated with several mechanical properties including yield and ultimate tensile strength, elongation, and notched tensile strength (Kt = 6.3). The yield and tensite strengths generally increased with increasing martensite content, but ) no correlation was observed between elongation and martensite content. The notched/un-notched ratio (Kt = 6.3) at -423°F exhibited a maximun at about 65 pct martensite. Attempts to correlate austenite stability (as measured by the Ms temperature) with both martensite content, and incremental increases in martensite content resulting from tensile testing were generally unsuccessful because the variable of cold Work could not be separated. As a result, only a general correlation of these variables was obtained. A (111) pole figure of austenite in a cold-rolled 301 stainless steel mas prepared an attempt to further investigate directionality as exhibited by mechan- ical properties. Yield strength was correlated with this pole figure, but ultimate and notched tensile strengths were not. 1 HE literature contains a wide variety of references pertaining to the austenite to martensite solid state phase transformation in austenitic stainless steel, see references 1 through 18. Discussions of the crystallography and kinetics of this transformation are given in references 2, 7, 8, 'and 11, while further discussions of austenite stability and hardening mechanisms are given in references 1, 3, and 10. Two important processing variables involved in this reaction are the degree of cold work and the temperature at which cold work is performed, and the nature of these effects are outlined in references 4 and 12, and 15 and 16 respectively. The latter two references deal primarily with cold rolling conducted at subzero temperatures. The effect of chemical composition on the austenite to martensite reaction was treated in detail by Eichelman and Hull, and they determined an equation for the M, temperature of this type of steel which included the effects of nickel, chromium, carbon, nitrogen, silicon, and manganese.5 Previous studies have shown that austenitic stainless steels of AISI types 301, 302, 304ELC, 310, and 301-N exhibit mechanical properties between 78" and -423"~ that are microstructure dependent, and the dependence of this microstructure on chemistry, temperature, tensile stress, and prior deformation is in general accord with previous investigations which have measured the effects of each of these variables on structure, or more specifically, the austenite-to-martensite transformation.lg The behavior of fully stable alloys, such as ATST type 310, at cryogenic temperatures is

Jan 1, 1962

By A. E. Perkins

fractured ingot will show a complete diagonal structure. If the temperature is lowered, the equi-axed zone is larger and the pine tree growth is less. A small ingot, though, is prone to this columnar structure, even though the casting temperature is kept down. The speed of cooling after solidification is slower in the hollow ingot than in the solid, therefore the granulation, which takes place after the crystals have formed from the liquid metal, is apt to produce larger grains than in the solid ingot. This is not a detriment. The original crystal size is of much more importance. The size of the grains due to this granulation constantly varies in response to mechanical work and subsequent heat treatment. Surface weakness, as shown by seaming and checking on the exterior when working or by cracking in hardening, is in some measure the result of the surface stresses of the ingot when cast. The temperature of the ingot mold or speed of freezing of the exterior of the ingot is of great importance and much seamy and cracked steel is caused by a too quickly cooled skin. Ingots that are turned or milled all over will not seam very much in working, in fact they are practically free from seams, whereas an ingot rolled without a preliminary machining operation is full of seams and cracks, some long and some short. It has been said that these short seams and cracks in the surface of the bar are the result of elongated depressions in the surface of the ingot and that if an ingot is cast very smooth it should give fewer surface defects. This is partly true, but an ingot cast very hot and having a very smooth skin will crack and seam very badly in the mill, as will the surface of an ingot poured very cold, which is very crinkly, because of the low temperature of the metal. The cause of the cracks and seams in one is not the cause of the cracks and seams in the other. The tube method of hollow drill steel manufacture is superior to the older method because of: (1) Greater freedom from external and internal straining; (2) the inherent small crystal size; (3) absence of harmful segregation, resulting in weakness of wall of hole. (4) less liability for steel to crack in inside of hole during forging or hardening; (5) toughening effect, arising from mild steel wall of hole limiting the intense hardening on quenching. Breakage and Heat Treatment or Rock-drill Steel By A. E. Perkins, Pittsburgh, Pa. Actual figures showing the superiority of one section or another are difficult to obtain. In a given size of steel, however, the cruciform sec -tion shows the most breakage; it is inherently a weak section. It contains less steel per foot than any other standard section of the same size and when rolled the wings tend to cool off setting up cooling stresses.

Jan 1, 1922

By O. Katz, A. Spilners, M. Simnad

THE oxidation of titanium has been studied in recent years by a number of investigators.'-W ne of the most important aspects of the reaction of titanium with oxygen, which has been observed by several of the investigators, is the extremely high solubility of oxygen in the metal. A detailed study of the Ti-0 systeml% as shown that the solubility limit of oxygen in a-titanium is approximately 14.5 wt pct 0 from 800" to 1700°C. The unusual oxidation behavior of titanium may be ascribed to this phenomenon, since part of the oxygen uptake enters into solid solution in the metal and only a part is available for scale formation. A more complete study of the kinetics of oxidation of titanium would require information on the influence of oxygen in solution in the metal upon the oxidation rate. Such information would also be of interest in regard to the theory of oxidation of metals. The authors have measured the oxidation rates of oxygen-saturated and oxygen-free titanium between 800" and 1200°C and have found appreciable differences in their rates of oxygen uptake and in the activation energies. The experimental set-up was similar to that described in a previous paper.u The course of oxida-dation was followed continously by attaching the specimen to a sensitive spring balance and observing the extension of the spring with time of oxidation. High purity electrolytic oxygen was used and precautions were taken to purify the gases employed. The specimens were prepared from iodide titanium and were ground on metallographic polishing papers to a fineness of 4/0. The furnace was first evacuated and filled with argon; the specimen was lowered into the furnace, which was then evacuated and filled with oxygen, after the specimen had reached the oxidation temperature in argon. The samples were presaturated with oxygen by means of the following procedure: Each specimen was first heated in oxygen at 1200°C until its increase in weight indicated 13.8 pct by weight uptake of oxygen. The furnace was then immediately evacuated and filled with purified argon five times. The specimen was heated in the argon atmosphere for a period of 72 hr at 1200°C in order to saturate the metal with oxygen by complete absorption of the scale. The temperature was lowered to the oxidation temperature, the furnace was evacuated and filled with oxygen, and the oxidation rate was followed by noting the extension of the spring balance. No oxide film was visible on these specimens at the start of the oxidation experiments. Titanium samples in the oxygen-free as-received condition were also oxidized and their oxidation rates measured. The results are shown in Figs. 1, 2, and 3, which indicate that the initial rates of uptake of oxygen by both the oxygen-saturated and oxygen-free specimens are parabolic at these temperatures. The rate of uptake of oxygen is, however, appreciably lower

Jan 1, 1956

By Pierre Auger, Benoit Levesque, Richard Martel, Guy Sanfacon, Louis-Charles Boutin, Marc-Andre Lavigne, Patrick Brousseau, Luc Trepanier, Louise Galarneau

Explosives used for blasting operations in civil engineering works, like construction of piping systems under roads, of pools, of houses and buildings can generate large volumes of carbon monoxide (CO). Carbon monoxide can represent between 1 to 3% of the gases produced by usual explosives. The production of 10 to 24 1 of CO by kg of explosives blasted is theoretically anticipated. CO is a gas that has no characteristic colour, smell or taste. It can migrate in the fractured rock of the blasted areas on fair distances, and then infiltrate in closed spaces like sewage systems, manholes, but also in basements of houses. In Quebec, in the last 10 years, 7 people were poisoned by CO in their houses to the point that they had to be treated in hyperbaric chambers. In the USA, many accidents have occurred and one worker died in a manhole. From the review of intoxication cases, the distance of migration of CO may reach 55 m if a conduit is present underground. Underground conduits broken by blast, tilling around underground conduits in road or house trenches, or fractured rock created by blasts between houses or between a house and a road, are the different pathways identified in the Quebec incidents. Field tests were performed at sites close to CO intoxication locations in order to evaluate the behaviour of CO in fractured rock following blasting operations. From these tests, the following conclusions may be drawn. The structural geology of the rock formation (schistosity, family of joints, fractures) plays a role in the direction and extent of gas migration in fractures generated when rock is blasted with explosives. The confinement of the rock can affect the quantity of gas migrating in the fractured rock. Significant concentrations of CO may persist in the fractured rock even 7 days after a blast. Advection is the initial mechanism of CO migration in fractured rock generated by the blasts or naturally occurring before the blast. The distance of migration with this mechanism is short (5 to 8 m). In trenches of muck (equivalent to tills under roads) the distance of CO migration by advection is 12 to 20 m. In the 2 to 3 days following the blasts, further CO migration up to 15 m is made by diffusion in the induced fractures, or up to 30 m in the trenches. To minimise CO production and migration in the fractured rock during blasting, special procedures can be applied by the blasting contractors: (1) do the complete excavation of the overburden before drilling and put mats for each blast; (2) do the complete excavation of the muck after each blast; (3) the last blasting should be the one for the private services entrance of the house and the excavation of the muck should begin at the junction between the road trench and the services entrance trench. Another technique consists in the pumping of interstitial air in the muck with a vacuum truck immediately after each blast.

Jan 1, 2002

By Roy J. Helfinstine, Gilbert H. Cady

One of the most urgent needs in the field of coal combustion is the ability to predict the performance of a coal from knowledge gained from small-scale tests. Numerous types of analyses and tests are conducted on coal, including the proximate and ultimate analyses, heating value, varieties of sulphur, ash analysis, ash fusion temperatures, free swelling index, petrographic analysis, and Gieseler. plasticity. The Illinois Geological Survey has made an intensive study of the value of these tests for judging the performance characteristics of the coal as actually utilized. To date, three reports1,2,3 have been issued, and another is being prepared. The present paper briefly summarizes some of the results found in the studies. Equipment The equipment used for the tests is described in some detail in the above mentioned publications. Briefly, it consisted of a domestic stoker, cast-iron boiler, and water-cooled heat exchanger, which were operated as a forced circulation hot-water system. The entire unit was mounted on scales. Instruments for recording the per- formance of the coal included a hot-water meter to indicate the quantity of water flowing through the boiler; a two-pen, mercury-actuated thermometer to record the temperatures of the water entering and leaving the boiler; a chemical-type meter to record the percentage of CO2 in the stack gases; a pressure gauge to record the static pressure in the stoker-air duct; and a multipoint potentiometer to record the temperatures in the stack and room, and the opacity of the stack gases. A 16 mm motion picture camera was available for taking pictures of either the fuel bed or the scale dial. Procedure The tests described in this report can be conveniently divided into three phases. For the first phase, a load of 5 or 6 tons of unwashed coal (usually screenings) was obtained from each of fourteen Illinois mines. These coals were all crushed and screened to a size of 35 in. (square hole) by 8 mesh. Combustion tests were made on a part of this raw, screened coal. Another portion was passed over a concentrating table to reduce the ash to what was considered to be commercially reasonable, and combustion tests were made on this upgraded coal. The ash content of the remaining coal was reduced to a still lower figure by passing it over the washing table with greater reject, or by standard float-sink procedure. This cleaned fraction was then tested in the stoker-boiler unit. Thus three coals with varying ash content were tested from each mine. For the second phase reported here, coals were obtained from four Arkansas mines and tested as received and after passing over the concentrating table with a normal reject. One Illinois coal was tested in the "washed" condition only. The third phase included tests on commercially prepared stoker coals

Jan 1, 1950

By David M. Parkes, Mike Fisher

In 1947, coal production in Western Canada was 10 Mt (11 million st) which predominantly came from 120 underground mines. By 1968 production was down to 1 Mt (1.1 million st). Production in 1981 was back up to 29 Mt (32 million st), but only 1.5 Mt (1.7 million st) of the total was from underground mines. In fact, because of competition from open-pit mines, only two underground mines were operating in 1981-the McIntyre and B.C. Coal mines. While several other underground mines are being planned, this article will deal with only the two in production. Geology The geology in Western Canada is frequently complicated; hence only more regular deposits are mined. Coal is generally thick, often steeply pitched and usually soft or well cleated. The 5000-km (3,100-mile) coal belt that runs from the Yukon to Mexico is well developed in Western Canada. It is divided into two distinct regions-the Rocky Mountain Lower Cretaceous coals and the mostly uniform level Upper Cretaceous coals of the plains area to the east. Isolated pockets of coal occur at Vancouver Island and Hat Creek. Hat Creek's massive 300-m (984-ft) thick geologically recent lignite deposit is 250 km (155 miles) inland. Geography The infrastructure of Western Canada is shown in Fig. 1 along with the locations of the two mines. Both rely on exports of premium low volatile metallurgical coal mostly to the Pacific Rim. The rail distance from both mines to port is 1100 km (684 miles)-accounting for a significant portion of the sales cost of the coal. Mining Methods Two main mining methods are used-room and pillar and hydraulic. Longwall has been attempted and is being considered again. In the past, a simple hand method used gravity to help mine most of the coal on steep pitches. Levels were driven in the seam, rising slightly to assist in draining. On the advance, mining would start at the upper level, usually from adits driven in the side of the mountain. Rooms or angles were driven up 100 m (328 ft) at a suitable angle for the coal to flow down. On retreating out of the level, the pillars would be taken, starting from the highest point, using batteries to stop coal or rock flushing down the room. Batteries consisted of three to six stout wooden posts set at an angle to resist material flow. Coal loading was directly from chutes at the bottom of raises into 1- to 2-t mine cars hauled by diesel or battery locomotives. Often output was restricted by the supply of empty cars. Mechanization began in the 1960s but was difficult in these conditions. Mines such as Canmore and Coleman continued operation into the 1970s but closed because steep seams, high methane emissions, and faulting made mining difficult. McIntyre Mines The McIntyre mines are located 360 km (224 miles) west of Edmonton, Alberta, in the foothills of the

Jan 4, 1983

By Virginia Lawson, Adam Price

The rougher flotation stage is used to reject liberated gangue at a coarse size. Base and precious metals due to their high specific gravity tend to recycle in the grinding circuit and report to the flotation circuit at a finer size than the gangue from which it is being separated. Generally liberation levels of greater than 50% are achieved in roughing and thus a portion of the feed can be immediately upgraded to final concentrate. The proportion increasing as the amount of liberated mineral in the feed increases. Scalping out concentrate from the rougher can reduce the size of the subsequent regrind and cleaner circuits. Examples of alternative process design for circuits demonstrates that you don’t need large volumes to effectively recover all of the minerals – you need smart flotation. Several case studies will be presented where circuits can be significantly reduced in size by designing smarter – not larger. This paper describes case studies to test the amenability of the Jameson Cell to produce final concentrate from rougher feed. Metallurgical performance and plant design implications are discussed. Froth flotation has been used extensively for mineral concentration for over 100 years. Although the flotation cell design has changed, it has generally simply increased in capacity rather than increase in efficiency. Cell capacities of over 600 m3 are now being used in rougher flotation particularly for the lower grade ores where the amount of surface area for flotation is deemed less significant than the required volume. The plant design volume in rougher flotation comes from laboratory batch flotation tests and/or pilot plants. The volume required for the plant is calculated based on the laboratory residence time multiplied by a scale- up factor. Examples of laboratory to plant scale-up design are given in Dunne et al. (2010 part 2). The factors are usually between 2 and 3. A laboratory residence time of 10 minutes would equate to a plant design residence time of 25 minutes using a 2.5 factor. The volume of the required flotation cells can then be calculated using the design tonnage and volumetric throughput. This scale-up factor takes into account the differences in performance between batch flotation efficiency and full scale plant. The differences are likely related to inefficiencies of bubble particle contact in large cells as well as froth recovery, which is essentially complete in a batch cell. Flotation cells, such as the Jameson Cells, are smarter designs that improve bubble/particle collection, bubble/pulp separation efficiency and operate with high froth recovery rather than simply getting bigger. The Jameson Cell was designed on the principal of improving the collection efficiency of fine particles by separating the contact zone from the separation zone and having the bubbles generated by the plunging jet shearing and entraining air. This means the method of bubble generation is not impacted by the wear of mechanical equipment as it the case in other flotation devices. Jameson Cells make bubbles less than 600 micron in size under most operating conditions (Ding, 2016).

Jan 1, 2019

By Leonard Bernstein, Harry Bartholomew

MANY of the properties of metals and alloys are structure dependent. Not the least of these is the grain structure. For example, in producing alloy bonds between silicon or germanium and gold-alloy foil at temperatures above their eutectics (which are, respectively, at 370" and 356°C) but below the melting points of the individual constituents, it was found that wetting was more complete with a fine-grain structure than with a coarse-grain structure. In cases where the assemblies were held together for long periods of time at elevated temperatures, wetting was not as critically dependent on the grain structure as was the case for shorter periods of time. When liquid phase forms, it first forms in the grain boundaries due to the fact that most of the impurities are concentrated at the grain boundaries and that most impurities tend to lower the melting point of the gold alloy somewhat. In addition, grain boundaries are at high-surface free-energy levels and are thus in a more favorable condition to initiate good wetting. A fine-grain structure will have more grain boundaries per unit area than coarse-grain structure thereby making wetting somewhat easier. To make the grain structure of an alloy sys- tem visible, a satisfactory method for differentiating between the different grains of which the system is composed must be employed. For the alloy bonding system previously described, the grain structure on the foil surface is of primary interest. Conventionally used etchants for the gold-alloy system will not adequately delineate grain boundaries in their structures. For this reason an alternate method is proposed for those cases where such a determination is necessary. Foil between 0.0005 and 0.005 in. thick has been used in this determination. The equipment used in this investigation was described in a previous publication.' Essentially, it employs a resistance-type filament which can be heated up very rapidly and cooled very rapidly. A thin refractory material is placed on top of the heating element. This refractory material may be molybdenum, tantalum, tungsten, or even ceramic but its thickness should be less than 0.010 in. to minimize the temperature gradient across it. The gold-alloy foil is placed on top of this refractory material. The temperature of the heating filament is raised to 900°C and cooled down to below 100°C in 5 to 10 sec. During this cycle, the foil can be observed to plastically deform due to its own weight. With no further chemical etching, this assembly is then viewed under a standard metallurgical microscope and the grain size measured. Fig. 1 is a Au-1 pct Sb alloy at 400X magnification and Fig. 2 is a Au-1 pct Ga alloy at 130X. This technique has proved to be an adequate means for controlling processing conditions in producing large quantities of gold alloy foil.

Jan 1, 1963

Abbott, Argyle Campbell 1209 Sherwin Ave., Chicago, Ill. Almstrom, Adne A., Student, Met. Engrg., Washington State College Pullman, Wash. Ankudinoff, N., School of Mines, Univ. of Utah Salt Lake Cite, Utah. Apter, Robert F., Student, McGill Univ Montreal, Que., Canada. Bailey, Donald G 798 Kensington Rd., Los Angeles, Cal. Baker, Kenneth B., Met., Aluminum Co. of America Pittsburgh, Pa. Barger, Thomas, Student, Univ. of North Dakota Grand Forks, N. D. Bartholomees, Geo. C Box 304, Bonne Terre, Mo. Bates, Walter, Student, Univ. of Utah, School of Mines Salt Lake Cite, Utah. Bobb, Russell V 16 College Ave., Houghton, Mich. Boswell, John C., Stripping Asst., Fairbanks Exploration Co Fairbanks, Alaska. Brandenburg, E. C., Student, Univ. of Kentucky 112 Church St., Lexington, Ky. Brannen, John, Chemist, Assay Office, Consolidated Min. & Smelt. Co. of Canada, Ltd., Tadanac, B. C., Canada. Brooker, Edgar, Jr., Student, Ferrous Met., Stanford Univ., Stanford University, Cal. Brooking, Leslie E 738 Lakoma, Norman, Okla. Clark, Martin Lester, Student, Missouri School of Mines Rolla, Mo. Collins, Edward D., Jr Box 1597, Ribbing, Minn. Comeau, Faye Bernard, Student, New Mexico State School of Mines Socorro, N. M. Conder, Henry S., Student, Pet. Engrg., College of Min., Univ. of California, Berkeley, Cal. Cotton, John Francis, Metallographist, Stealton Plant, Bethlehem Steel Corp., Steelton. Pa. Dahlgren, Elmer George, Student, Univ. of Wisconsin, Tripp Hall, Madison, Wis. Dale, James L., Student, Colorado School of Mines Golden, Colo. Dashti, Mirza Abdullah Khan, Graduate Student, Columbia School of Mines, Engrg. Chemistry, New York, N. Y. DeHuff, Gilbert L., Jr., Student, Lehigh Univ Bethlehem, Pa. de Romana, Enrique P., Jr. Engr Casilla 22, Arequipa, Peru. Oresbach, Charles Howard, Geol., Gulf Production Co Box 1428, Amarillo, Tex. Evens, Howard G., Student, Michigan College of Min. & Tech Houghton. Mich. Fassler, Glen Emerson Rodman, Inspiration, Ariz. Feight, Charles Donald, Foreman, Clairton By-Product Coke Wks., Carnegie Steel Co., 319 5th St., Wilson, Pa. Floe, Carl F., Student, State College of Washington Pullman, Wash. Fuentes, Vicente, Student, Pet. Engrg., Univ. of Oklahoma, School of Mines, Norman, Okla. Fuhrman, Carl F., Student, New Mexico School of Mines Socorro, N. M. Goergen, Gerald G., Cerro de Pasco Copper Corp Casapalca, Peru. Grove, Hubert A., Training Course, U. S. Aluminum Co New Kensington, Pa. Gastetter, Robert E 1010 N. Cherry Ave., Tucson, Ariz. Halbouty, M. T., Student, Texas Agriculturcl & Mech. College, Box 254, Faculty Exchange, College Station, Tex. Halfast, Edgar 710 E. Okmulgee Ave., Muskogee, Okla. Hall, Robert Lee 121 58th St., Niagara Falls, N. Y. Hanson, Rolf H., Student, Colorado School of Mines Golden, Colo. Hillyer. Shaler G., Student, New Mexico School of Mines Socorro, N. M. Hoertel, Frederick W., Met., R. W. Hunt Co., Engrs St. Louis, Mo. Holtz, Henry G., Student, Ohio State Univ Columbus, Ohio. Hon, Richard J Box 721, Miami, Ariz. Huffman, James O'Neil, Student, College of Min., Univ. of California Berkeley, Cal.

Jan 1, 1929

By L. Max Scott

[INTRODUCTION] The significance of gamma exposure to uranium miners has been recognized only in the last few years. Most ore deposits which have been underground mined, were 1% or less U308. Full-time mining of this grade ore can result in exposure exceeding 1 Rem per year. Several companies in Saskatchewan are planning to mine recently discovered ore bodies which contain ore pods in excess of 10% U308. The purpose of this paper is to present dose data which can be used to estimate gamma exposure from high-grade ore deposits, and to present mining techniques which will minimize miner exposure. [Dose Estimates] When one has a point source of radiation, the conventional laws of isotropic emission and the fall of dose with square of the distance, make dose calculations relatively easy. However, if one has an area source of infinite thickness, the assumption must be made that a uniform gamma flux is being emitted from all faces of the source (Figure 1). Dose from gamma emission would fall off only due to air attenuation up to such a distance at which the source can be considered to be a point source (usually 10 times the largest dimension of the source). For short distances, such as to the cab of ore-moving equipment, the dose reduction would be negligible. Assumptions: 1) Radium and gamma-emitting daughters are in secular equilibrium with uranium. 2) All gamma radiation is emitted from radium and its daughters. 3) Average gamma energy of 0.8 MeV. 4) Uranium ore has a linear attenuation coefficient of 0.212 cm-1, similar to that of Si02(sand). 5) Uranium ore density of 3.0 gm/cm3. 6) One curie of radium emits 6.84 x 1010 photon/sec(l). 7) One percent uranium ore contains 1 x 10-8 Ci/cm3 of radium. 8) Each MeV of 0.8 MeV photon results in a dose of 2.10[ ]10-3 mRem/hr. per MeV/cm2 -sec. 1) 9) Minimum ore depth 100 cm. 10) Linear attenuation coefficient instant for radium gamma of 0.374 cm -1 [ ] The photon flux from a radioactive slab is computed as follows: [Flux = photon/cm3 3(1) 2u S] [where, us is the linear absorption coefficient of s the source (uranium ore).] Thus, 1% uranium ore, in the form of an infinite slab, emits 1613 photon/cm2-sec, which results in a dose rate of 2.71 mRem/hr. Since the dose from an infinite slab is the same as a dose from an infinite radius right circular cylinder, the only correction needed from the above dose rate is for finite radius and point of dose above the end of the cylinder (Figure 2) [Dose = mRem/hr. {E2(bl) - E2(b3) - [(E2(b1sec[E] -E2(b3sec[E]]cos[O](1)]

Jan 1, 1981

By J. H. Kittel

THE properties of the metal molybdenum have been the subject of recent investigations1,2 = which have included studies of the changes that occur in the metal when it is annealed after having been cold worked. Most of the data which have been reported have pertained to only one aspect of annealing, that is, the conditions under which recrys-tallization has been observecl. The process of recovery which may precede recrystallization docs not appear to have been investigated, and it is the purpose of this note to report observations on the recovery process in cold-worked molybdenum as determined by changes ill the thermoelectric power which occur as the metal is annealed. The measurements were made on commercially pure molybdenum wires which, after annealing at 1035°C in dried hydrogen for 1 hr, were swaged to give a series of reductions up to a maximum of 84.5 pct. The variously cold-worked wires were then annea1c.d in dried hydrogen for periods of 16 hr at progrc:ssively highel. temperatures from 200" to 1000 c with a final annealing treatment of 70 hr at. 1100°C. After each annealing treatment the ther.ma1 emf's were determined against a length of original unwo~,ked material that had received the initial hydrogen anrit,aling treatment. The difference in temperature between the hot and cold junctions was maintained at about 75'C. and. the thermal emf's were measured with a potentiometer that could be read. to 0.01 microvolt. Fig. 1 shows the thermoelectric power. (dE/dT) in microvolts per C between annclalcd and cold-worked molybdenum as a function of the degree of cold work. The effect of annealing is shown in Fig. 2. The general shape of the curves in both figures is similar to results which have been obtained for copper." It was observed that after annealing at temperatures as low as 200°C recovery was well under way. The rapid changes irt thermoelectric power observed near 900 C are probably associated with recrystallization, since such a relation has been observed with copper. Beyond 900°C the thermal emf's changed in polarity. indicating that from this point or1 the specimens were more highly annealed than the standard. Annealing treatments could not be given at temperatures higher than 1100°C due to limitations of the furnace, and it is evident that, since equilibrium values of thermoelectric powers were far from having been attained, temperatures considerably higher are required to fully anneal molybdenum. References Robert M. Parke: Molybdenum, A New High-Temperature Material. Metal Progress (1951) 60, pp. 81-96. * R. A. Long, K. C. Dike, and H. R. Bear: Some Properties of High-Purity Sintered Wrought Molybdenum Metal at Temperatures up to 2400 °F. NACA TN 2319 (March 1951): G. W. Brindley: A Thern-io-Electric Study of the Cold-Rolling and Heat-Treatment of Copper, Report of a Conference on the Strength of Solids. Physical Society, London (1948) pp. 95-106.

Jan 1, 1953

By R. V. Smith, M. H. Cullender

Rigorous equations for calculating subsurface pres.tures in flowing and static gas wells, and pressures along horizontal pipelines are presented in this paper. These general equations, based on the mechanical encrgy balance, contain no assumptions regarding temperature and can be used with either straight-line or curved temperature gradients. The friction factors recommended in this report are based on an absolute roughness of 0.0006 in. Flow is always considered to he turbulent. Although these general flow equations are solved by numerical means, the methods are as convenient as many of those used today to calculate pressures in gas wells and pipelines. These numerical methods are illustrated for flowing and static columns of gas in wells and for flow in pipelines. The authors believe that the use of these methods will result in more accurately calculated pressures in gas wells and pipelines and that [he methods are more flexible than any in use at the present time. Also, the methods are easily adaptable to automatic computers. INTRODUCTION The problem of calculating subsurface pressures in gas wells and along flowing pipelines has been studied by many investigators with the results that two widely-used methods have been presented in technical literature. One approximates conditions in a gas well or pipeline by assuming that temperature and compressi- bility are constants tor the entire gas column. The other method assumes that temperature is constant at some average value but permits compressibility to vary with pressure at the constant temperature. While these approximations may be justified for relatively shallow and low-pressure wells, they become more and more unrealistic as the calculations are applied to the deeper high-pressure wells. In wells or pipelines where there is an appreciable change in temperature between the inlet and outlet ends of the flow string, it is equally as important to consider the change in compressibility with temperature as it is to consider the change with pressure. Consequently, the methods presented in this paper were developed to provide a workable procedure for calculating pressures in gas wells and pipelines which makes no assumptions regarding either temperature or compressibility. The friction factors recommended for use with the equations were the result of study of published literature and extensive experimental work with gas wells. The F,. values in Appendix IV which contain the friction factor are equivalent to an absolute roughness of 0.0006 in. as calculated by the Nikuradse equation. The absolute roughness of 0.0006 in. was previously published by Cullender and Binckley2 for gas wells; it is in agreement with 0.00065 in. of Smith, Dewees, and Williams1 for gas wells and 0.00065 in. of Baker and Swerdloff4 for pipelines; it is in reasonable agreement with 0.0007 in. reported by the Bureau of Mines and the American Gas Assn.5 for large transmission lines. However, any type of friction factors may be used in the equations by a change in the F, values in Appendix [V].

Jan 1, 1957

By John E. Dutrizac, Tzong T. Chen

"Laboratory experiments were undertaken to ascertain the behaviour of various ""impurity"" species during the precipitation of hematite directly from ferric sulphate solutions at 225°C and in the presence of 40 or 50 g/L of hematite seed. Copper, nickel and zinc are not significantly incorporated in hematite precipitates. Although the nickel contents increase to 0.22% as the Ni concentration increases to 60 g/L, the solid solution Ni contents are low relative to the bulk Ni contents. Thus, part of the Ni may be present as tiny particles of poorly soluble NiS04.H20. Increasing germanium concentrations to 2,000 mg/L result in products containing up to 0.59% Ge. The Ge occurs either in solid solution in the hematite or is adsorbed on the nanometer-size hematite crystallites. Increasing As(V) concentrations to 10 g/L increase the AsO4 contents of the products up to 16%. However, the products consist of hematite together with As04-containing Fe(SO4)(OH) and an unknown Fe-AsO4 species. All the hematite products contain - 1% SO4 which appears to be adsorbed on the nanometer-size hematite crystallites.IntroductionDissolved iron is a common impurity in most hydrometallurgical processing circuits and, generally, the iron must be eliminated prior to the subsequent recovery of the sought after metals. Simple neutralization of iron-rich solutions yields slurries which are difficult to thicken and filter. To address these concerns, the jarosite, goethite and hematite processes have been developed, and these technologies are able to reject the dissolved iron in a readily filterable form [1]. Despite its higher cost, the hematite process offers several potential advantages. Because hematite is an iron-rich, high-density compound, its formation would reduce the mass of residue and the size of a residue storage pond. Hematite is the thermodynamically stable form of iron under ambient conditions, and this stability would facilitate the impoundment of hematite precipitates. Furthermore, hematite has some market potential for cement manufacture and even for iron making [2]. In all the commercial hematite processes, the dissolved ferric ions are first reduced to the ferrous state. Subsequently, the pH of the solution is adjusted and the solution is then oxidized. The ferric ions gradually generated during solution oxidation hydrolyze and precipitate mostly as hematite, although commercial hematite products also contain trace or minor amounts of jarosite-type compounds. However, there are potential advantages in precipitating hematite directly from ferric ion solutions, as the reduction and subsequent oxidation of the iron are not required. Although the precipitation of hematite directly from relatively concentrated ferric sulphate solutions (up to 0.5 M Fe(SO4)is) was recently demonstrated [3], the behaviour of impurity species during hematite precipitation is largely unknown. To address this issue, systematic laboratory experiments were undertaken to ascertain the behaviour of a number of ""impurity"" species during the precipitation of hematite directly from ferric sulphate solutions at 225°C, and the results are presented in this paper."

Jan 1, 2012

By Jason Herron, Joseph Brown

The common couplings for joining sections of helical piles with an outside diameter of 6.625 inches and 8.625 inches are square on square, flange to flange, external pipe sleeve, internal pipe sleeve, and welded joint. Each coupling type has advantages and disadvantages in efficiency, capacity, installation, and economics of the installed pile. All these coupling types have been used in the helical pile market in North America for various applications and designs. Many of these applications require on-site testing, but test records are not typically shared with the helical pile industry. Due to the lack of information available for square on square and flange to flange couplings, testing was done to compare the efficiency of installation, load to deflection response, and torque to capacity ratios of external pipe sleeve, square on square, and flange to flange couplings. Field installations and tension tests were performed in non-cohesive (sand type) soil near Caseyville, IL and in cohesive (clay type) soil in Centralia, MO. The data collected during these tests included rotations per foot during installation, installation torque, applied pressure, and deflection. The results of this testing are analyzed and discussed within this paper. The results indicate that the square to square coupling is the most efficient of the designs tested, but this conclusion cannot be fully justified without further testing due to the variability in the data gathered and the small number of piles tested. INTRODUCTION A helical pile is typically assumed to be a medium-displacement pile when its outside diameter is between 4.5 inches and 8 inches and a high-displacement pile when its outside diameter is greater than 8 inches. A helical pile designer has multiple options when selecting a coupling type for medium- to high-displacement helical piles. Each type of coupling has advantages and disadvantages and must be analyzed on the bases of efficiency, capacity, installation, and economics of a project. Typical couplings that are chosen by designers include square on square, flange to flange, external pipe sleeve, internal pipe sleeve, and welded joint. A helical pile must transfer axial, lateral, bending moment, and torsional loads, and the coupling must be designed to effectively and efficiently transfer each type of load. During the installation of a helical pile, the rate of penetration and the disturbance of soil layers play a critical role in the efficiency and torque to capacity ratio (Kt) of the pile. It is common for a coupling to protrude beyond the outside diameter of the pipe section being used. This is necessary for the function of certain coupling types and can facilitate the connection of couplings. However, it causes increased soil disturbance and reduce the rate of penetration due to soil bearing against the protruding portion of the coupling. A helical pile designer typically will work to minimize the coupling protrusion and create a pile with optimal installation efficiency but may choose a design with lower installation efficiency if it offers sufficient economic advantage.

Jan 1, 2019

By E. Asselin, P. Sauve, H. Salomon-de-Friedberg

"Residues from the CESL Copper process were leached using the CESL Gold process. The kinetics of reaction were confirmed to be fast with over 50% of the gold consistently extracted within 2 minutes of leaching. Final gold extractions under the conditions tested here were between 80-90% obtained within 90 minutes. A model for the CESL Gold process was developed. This model is slightly different based on whether the copper concentrate leach residue was ground or unground prior to leaching: the gold leaching rate constant is approximately twice as fast for the ground samples. Unground residues had a P80 of 13 ?m whereas ground residues had a P80 of 10 ?m. The model predicts a 43% decrease in leaching time for ground residue versus unground residue to achieve the same extraction under otherwise identical conditions. The model was more accurate for the ground samples. INTRODUCTIONTeck's CESL Process (Jones, 1996) was developed to provide an economically attractive alternative to traditional copper smelting. It has since evolved to also deal with copper concentrates containing valuable impurities such as nickel and undesirable ones, such as arsenic, while reducing downstream costs. Effective recovery of gold and silver is essential to the economics of the CESL Process. The Brook Hunt 2007 copper concentrate assay data (Wood-Mackenzie, 2007) were examined to determine the ratio of gold plus silver value relative to the contained copper. As indicated in Figure 1 below, nearly two-thirds of concentrates have 10% or more of their total value tied up in their precious metals content.CESL has extensively examined the recovery of gold and silver at bench, pilot and demonstration scale, which has culminated in the present CESL Gold Process (Barr et al., 2007). The process targets the precious metal contained in the leach residues produced in the copper portion of the flowsheet. This residue has two primary components which influence downstream processing. The major iron-rich fraction (-75 wt%) of the residue is a mixture of various forms of iron oxide precipitates including hematite and jarosite (Sahu and Asselin, 2011). The iron fraction ranges from amorphous to crystalline in nature depending on the compound precipitated, upstream operating conditions and feed material composition. The lesser, but by no means less important, fraction is a mixture of elemental sulphur and unreacted sulphides (principally pyrite).Elemental sulphur formed during the oxidative leach of the copper concentrate reacts with cyanide to form thiocyanate (SCN-) making a traditional cyanide leach approach too costly. Since the SCN formation reaction strongly depends on contact time with the CESL residues (Po, 2007) short contact times were preferred. This factor guided the development work towards the eventual room temperature, pressure cyanidation process which results in near complete gold dissolution in 90 minutes. A recent CESL gold demonstration campaign treating 1.5 t/d of residue from the copper oxidation step achieved an average gold extraction of 91% (96% maximum)."

Jan 1, 2012

By B Smith, R Gelson

Springvale Colliery is an underground longwall mine situated approximately 10 km north-west of Lithgow in the western coalfields of New South Wales. Following a full-scale review of the mineÆs ventilation system in mid-2010, it was clear that the mine by early 2013 would become ventilation constrained as a result of the increasing size of the mine workings, the predicted worsening tailgate conditions and plans to introduce a fourth continuous miner unit. This realisation led Springvale to embark on a major upgrade of the mine ventilation system that would meet short-term and life-of-mine (LOM) requirements.Included in the upgrade project was the installation of a new high pressure centrifugal main exhaust fan that is capable of overcoming the high resistances imposed frequently on the circuit by restrictions in the tailgate airway. The new fan featuring a 2.5 MW motor is designed to be capable of operating at a quantity of 250 m3/s at a pressure of 7.9 kPa, a substantial improvement over the previous system, which had an operating duty of approximately 145 m3/s at 1.9 kPa. The new fan is of a centre hung design and features a backup diesel motor to maintain ventilation in the event of a power outage.Structural engineering analysis determined that under the operation of the new main fan the existing standard of ventilation control devices (VCDs) at Springvale would not be adequate to withstand the increase in differential pressures. As such, an upgrade of critical VCDs was undertaken utilising a novel high flexural strength (9 MPa) shotcrete and thin spray-on liner, which help serve to safeguard the structural integrity of the appliances and prevent increased leakage as roadway conditions at VCD sites deteriorate over time (a serious problem in many areas of the workings at Springvale).In addition to the main ventilation system upgrade, a separate upgrade focusing on improving development face ventilation was also undertaken. This upgrade included the purchase of two high capacity (22 m3/s) variable inlet vane (VIV) auxiliary fans to facilitate comfortable development face ventilation.The upgrade of the mine ventilation system at Springvale has been three years in the making, but has removed a major business risk in that it has afforded mine planning and operational flexibility (a fourth unit can now be introduced) for the LOM, and importantly provided unprecedented levels of comfort to operators at the working faces where previously ventilation and heat issues were frequently reported. The impact of this on the morale of the workforce, and hence their output, should not be underestimated.CITATION:Gelson, R and Smith, B, 2013. Life-of-mine ventilation system upgrade at Springvale Colliery - A case study, in Proceedings The Australian Mine Ventilation Conference , pp 79-86 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By Henrik Gudbrandsen, Geir Martin Haarberg, Ana Maria Martinez, Karen S. Osen, Sverre Rolseth

"The Several approaches were tried for developing an electrochemical route for producing high purity silicon from molten salts. K2SiF6 and metallurgical grade silicon were used as the source of silicon. Molten LiF-KF and pure KF with additions of K2SiF6 were used as electrolytes at temperatures from 550 - 900 oC. Silicon was cathodically deposited and anodically dissolved in the fluoride electrolytes. High electrorefining efficiency was obtained for many elements after electrolysis experiments at constant potential or constant current. The main challenge was to reduce the contents of boron and phosphorus. Another difficulty was to remove inclusions of solidified electrolyte. The lowest energy consumption was estimated to be ~ 3 kWh/kg Si, while the best current efficiency for Si deposition was found to be ~ 95 %.IntroductionSilicon is the most important material for photovoltaic power generation devices because of high feasible efficiency, low ecological impact and slow degradation. Development of a new and less expensive route for producing silicon of the desired quality is necessary in order to make solar cells competitive. Electrochemical routes may be an alternative possibility, and many laboratory studies have been reported over the past few decades. Most of the studies have been performed in molten salts, chloride or fluoride, at moderate temperatures so that solid silicon was deposited at the cathode. The production of high purity silicon has shown an annual increase of about 25 %, due to the increasing demand of solar grade silicon (SoG-Si). For future increased use of solar energy it is important to develop alternative, more energy efficient and cheaper processes for SoG-Si production compared to the Siemens process.Common silicon sources have been SiO2 and K2SiF6 for electrowinning and MG-Si in the case of electrorefining. A serious challenge is to find an oxygen evolving anode for electrolytes containing dissolved oxides. Especially at moderate to high temperatures and in molten fluoride electrolytes this is a difficult task.Many experimental results have been reported in molten chloride electrolytes. Nohira et al. [1-3] and Chen et al. [4] cathodically deposited silicon by electrodeoxidation of solid SiO2 in molten CaCl2 by potentiostatic electrolysis. An alternatice approach is to use electrorefining of low-cost metallurgical grade silicon (MGSi) to produce SOG-Si. In previous studies [5], an MG-Si anode was used, and silicon was successfully electrodeposited on a molybdenum wire in a CaCl2 based melt. It was found that the anode was strongly passivated due to the precipitation of silicon. Also, some deposits detached from the molybdenum wire because of the poor adherence to the cathode.Considering the Cu-Si alloy which has a good segregation effect for some impurities, and should not passivate, could be a good candidate anode for electrorefining. [6]."

Jan 1, 2014

In Canada and other countries, legislation dealing with accurate factual disclosure for mining projects is being expanded to include more metallurgical design information. Grinding characteristics, metallurgical recovery of contained metal and environmental hazards associated with tailings disposal will be part of the Canadian package. This paper will examine only that portion of the plant design standard practice that relates to grinding ? in SAG mills and ball mills that grind ore from as received crushed material to final ground product that is fed to downstream plant recovery processes. Since grinding plants represent a significant portion of the overall concentrator cost, accountability in grinding design should no longer be left to chance, opinion, or hidden proprietary information. This is the 21st century. It is reasonable to expect that a factual basis will be needed so that a Qualified Person (QP) reviewing the grinding mill design can understand it. That will also involve how many and what quality of samples are taken to define the characteristics of the deposit, the accuracy of the tests used to measure ore hardness and the method by which the dimensions and power of the mills are determined. If these criteria are not met, the QP may have no choice but to call attention to the lack of convincing information regarding the grinding mill design for the project. At the same time clients are expected to demand getting the correct design information at a cheaper price, results in a more timely way, and accurate design to the point where the risk of tonnage shortfalls are reduced or eliminated. Tables will be presented to show how this can be done to current standards using existing technologies. At the same time it is acknowledged that current standards are ill-defined and that to achieve consensus among those involved in SAG mill design work, it may require that a consortium funded project be set up to establish what tests and methods for grinding design are suitable to meet the new standards for metallurgical disclosure. By running parallel design methods on several major projects over a period of 5 years, the process of how to design grinding circuits will be better understood because the projects will start at the design stage, include procurement of the mills and will end with a benchmark test on the plant. Selecting the samples and the best schedule for doing test work on core samples submitted and mill design, will be discussed in this paper. The issue of completing the grinding design in time for early procurement decisions is on the critical path for most major projects. Discussions between mining engineers, geologists and process engineers are needed early in the development cycle, at the scoping study level (SS), preliminary engineering assessment (PEA), prefeasibility study (PFS), and at the feasibility study stage (FS), to ensure that proper planning for design sample selection is done. Using adequate sampling and testing methods in a timely way allows early procurement of the grinding mills and minimizes project risk.

Jan 1, 2012