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By R. Le Houillier, M. Nagamori, A. J. Plumpton

"Asbestos tailings are a typical serpentine, hydrated magnesium silicate, consisting primarily of antigorite and a few per cent of unrecovered fibrous chrysotile. Serpentine has unique thermal properties in that dehydroxylating calcination at 600 to 700°C can yield an amorphous structure in which most of the magnesia is liberated from the original bond with silica and thereby rendered chemically active. The amount of activated magnesia can be quantified by selective leaching with a weak acid such as acetic acid. Many wet processes have been proposed to extract magnesium values from calcined or natural serpentine by means of various acids and their ammonium salts. However, the only commercial utilization of serpentine today is found in dry processes such as the production of fused magnesium phosphate fertilizer .The present paper reviews the physical chemistry of the dehydroxylation of serpentine and also various wet and dry processes in order to provide a general view regarding the chemical transformation of asbestos tailings.IntroductionAsbestos fibers occur as chrysotile, a serpentine polymorph. More than nine tenths of mined asbestos ores consist of massive non-fibrous serpentine, primarily the polymorph antigorite, which is discarded as tailings after fiber extraction. For many decades, Quebec has been a world leader in the production of asbestos fibers. As a result, colossal tonnages of serpentine have accumulated in the Eastern Townships'!'. Serpentine minerals are found in large deposit s throughout the world, and significant research efforts have been made, particularly in Japan and the USSR, mostly in attempts to extract magnesia or silica values therefrom.The chemical composition of serpentine (chrysotile and antigorite) may be ideally expressed as Mg3Si2Os(OH). which corresponds to 43.7% MgO and 43.3% SiO2. Natural serpentines often contain a few per cent of Ni-bearing magnetite and other impurities, and therefore the contents of MgO and Si02are both about 40% or lower.Serpentine (hydrated magnesium silicate) is virtually insoluble in weak acids. However, when dehydroxylated under controlled conditions, ' it decomposes to an amorphous state in which magnesia and silica are isolated and chemically active. The magnesia so activated is readily soluble in weak acids. This unique characteristic of calcined serpentine, first discovered in 1942,, has since attracted the attention of innumerable researchers, whence many processes have been proposed."

Jan 1, 1980

By Jean-Paul Drolet

"The ultimate aim of the Law of the Sea Conference, held under the auspices of the United Nations, is to ensure that all aspects of the use of the sea are covered by international law. The question facing the nations of the world can be stated as: "" Who will control and benefit from the use and the wealth of the sea?""The United Nations Conference on Law of the Sea - known as UNCLOS - which reconvened in New York during the latter part of August 1978 was the third such Conference in the UN's history. The concepts of sovereignty over the territorial sea and freedom on the high seas have remained fundamental to the international law o f the sea until our time. The legal regime based on these two concepts developed initially from state practice meant that the unilateral actions taken by one or more countries were eventually accepted and followed by others.It was not until 1958, during the first Geneva Conference on the Law of the Sea (pursued in a second Conference in 1960) that the Law of the Sea was codified and modified to a certain extent by the then independent nations of the world in the four multilateral conventions on territorial sea and contiguous zone, continental shelf, high seas, and fishing and conservation of living resources.A third Conference got under way many years later when the members of the UN met in New York at the end of 1973 for preliminary discussions . This meeting stemmed largely from the need to fill gaps in existing international legislation which did not cover the new uses of the sea. The Conference was subsequently reconvened a number of times - in Caracas (1974), in Geneva (1975) and twice in New York (1976) for what was called "" substantive negotiating sessions."" A sixth session was held in New York in 1977; a seventh, held in Geneva at the beginning of 1978, recessed and reconvened in New York at the end of August 1978. This is what is referred to, in international jargon, as the seventh session of the Third United Nations Conference on the Law o f the Sea. The following dates are the principal markers of a saga which began some twenty years ago with the purpose of regulating the use of the oceans."

Jan 1, 1981

By R. L. Mauger, P. E. Damon

Potassium-argon dating of the late Mesozoic and Cenozoic intermediate to acidic plutons and volcanic rocks of Arizona and northern Sonora demonstrates the existence of two distinct magmatic episodes. The earlier episode begins in late Campanian time and dies out before the middle Eocene. Following a quiescent period in middle and late Eocene time, magmatism increases in Oligocene time, becomes most intense at the Oligocene-Miocene boundary and dies out as the Pliocene is approached. Pliocene magmatism is primarily confined to extrusion of post-orogenic basalts. The late Mesozoic-early Cenozoic magmatic pulse is essentially confined within the limits of the classical Laramide orogeny (Laramie through Wasatch time). The mid-Tertiary pulse is contemporaneous with the mid-Tertiary Basin and Range orogeny. The Laramide provides an excellent specific example of Umbgrove's concept of the pulse of the earth and in general his concept appears to be meaningful. Copper porphyry mineralization is time-congruent with Laramide magmatism and there seems to be no doubt as to the existence of a genetic relationship between the two phenomena. Two origins are tenable: direct accumulation of the copper sulfides in the copper porphyry liquid magmatic environment; or, introduction of the mineral constituents entirely from a larger source magma when the host rock is already crystalline. Transgression and regression of the epicontinental seas provide an excellent indicator of orogeny and epeirogeny. There was an overall transgression of the epicontinental seas from early Triassic time through the Turonian epoch in early late-Cretaceous time. Thereafter, there is a relatively continuous regression of the seas until the present time. Both the earlier transgressive episode and the later retreat are intermpted by temporary but distinct reversals. These interruptions can be correlated with classical orogenies. There are six distinct orogenies so indicated in the Mesozoic-Cenozoic eras. These occur at approximately 35 to 40 m.y. intervals. The regression of the epicontinental seas following the Turonian epoch was contemporaneous with the deepening of the Pacific ocean basin over the Darwin Rise. As the Darwin Rise was collapsing the East Pacific Rise was being elevated. The Western United States, situated on the eastern flank of the East Pacific Rise, was upwarped and tilted toward the east. Epeirogeny is the primary consequence of processes that operate on a global scale. Orogeny is an exothennic process that is limited in space and is treated here as a secondary consequence of epeirogenic-initiated warping. The off-campus buildings of the University of Arizona Geochronology Laboratories rest on the volcanic rocks of Tumamoc Hill. From this hill, the observer may obtain a beautiful view of typical Basin and Range topography. The city of Tucson lies below, contained within the broad Santa Cruz basin and surrounded on all sides by characteristic NW-SE trending ranges towering as much as 7000 ft above the valley floor. Since the senior author arrived in Tucson eight years ago, a great deal more has been learned about the geochronology of this and neighboring basins and ranges. For example, the volcanic rocks underlying the laboratory at Tumamoc Hill were shown on the state geologic map as Quaternary basalt. A student of the senior author has demonstrated that the so-called basalts are in fact alkali-andesites (Halva, 1961), and we now know that these alkali-andesites or doreites (Mielke, 1965) are of early Miocene age rather than Quaternary age (Bikerman and Damon, 1965). Is it any wonder then that geologists disagree on the nature and time of Basin and Range volcanism, plutonism and tec-tonism? The plain fact is that few definitive Cenozoic fossil sites were available upon which to base a chronology and geologists were forced to guess the ages of rocks in order to achieve the semblance of a time classification. We have now made considerable progress in developing a K-Ar chronology for the volcanic and plutonic rocks of the Basin and Range (Bikerman, 1965; Damon and Bikerman, 1965; Damon et al, 1964 and 1965; Mauger et al, 1965). It is the purpose of this paper to point out the main features of this chronology and to discuss the relationship of magma-tism to orogeny-epeirogeny during post-Turonian time. We will also discuss probable relationships during

Jan 1, 1967

By R. W. Heckel, R. L. DeGregorio

The DeHoff method of determining the size distribution of ellipsoid-shaped, second-phase particles has been applied to the spheroidization of cementite in a eutectoid steel. The surface area of Precipitates determined from the various size fractions in the distribution was correlated with surface-area measurements based on the number of intersected interfaces on a random lest line. The precipitate particles were found to be oblate ellipsoids with an axial ratio of about 0.90. The size distributions were found to be log-normal. A method is proposed whereby the shrinkage of small particles and growth of large ones can be determined from the experimental data. The experimental data are compared to previously proposed mathematical models describing diffusion-controlled kinetics and various types of interface-controlled kinetics. The experimental growth and shrinkage rates are considerably slower than those predicted by diffusion-controlled kinetics. The best fit is obtained for a model describing interface-controlled kinetics limited by the rate of formation of cementite at the growing interfaces where the interfacial reaction rate is proportional to the solute thermodynamic activity gradient across the surface. THE subject of spheroidization of second-phase particles has been considered previously by other investigators. Livingston1 has studied the precipitation of the cobalt-rich phase from copper alloys containing 0.7 to 3.2 wt pet Co. His results indicate that the average particle size increases as the cube root of the heat treatment time in accord with diffusion-controlled kinetics.'-' Komatsu and rant' in their studies of the growth of SiO2 in a dispersion-strengthened copper-silica alloy found that the initial growth of SiO2 proceeded by diffusion-controlled growth and later stages were limited by interface-controlled growth. The values of the activation energy obtained for the interface- controlled process led them to the conclusion that the process was limited by the dissociation of SiO2 at the shrinking interface. The transition from diffusion-controlled growth to interface-controlled growth was characterized by a particle size which varied with the heat-treatment temperature. It is also interesting to note that the particle-size distribution as found by Komatsu and Grant exhibits log-normal behavior when replotted on log-probability coordinates. Dromsky, Lenel, and Ansell9 have observed the growth of Al2O3 particles having a mean free path of about 1.5 to 13 µ. Their photomicrographs indicate that the Al2O3 particle sizes were larger than those observed by Komatsu and Grant. Dromsky, Lenel, and Ansell concluded that the Al2O3 coarsened by an interface-controlled growth mechanism limited by the solution of A12O3 at the shrinking interfaces. Bannyh, Modin, and Modin10 have studied the spheroidization of a eutectoid steel (0.83 wt pet C). Their spheroidization (tempering) treatments were carried out in the range from 210o to 700°C for times between 1.5 sec and 20 hr. Measurements of mean particle size as a function of time indicated a cube root of time dependence of size at 700°C, in agreement with previous analyses of diffusion-controlled kinetics.2°7 At lower temperatures, the time dependence was less than the one-third power. It is important to note that, although mathematical models of growth* considering particle-size distribution have been available, measurements of only mean particle size have been carried out. DeHoff11 has presented a quantitative metallography technique which is applicable to the determination of the size distribution of ellipsoids of constant shape. This method is applicable to both oblate and prolate ellipsoids of all ratios of minor to major axis and is based upon an extension of Saltykov's analysis12 of the distribution of spheres of varying size. DeHoff's analysis is based upon measurements of the minor axis of the particles on a random plane of polish in a unit area. This method provides a measure of the number of ellipsoids in various size ranges per unit area. As pointed out by DeHoff, such measurements may be used to obtain surface area and total precipitate volume data. Thus, the accuracy of the distribution analysis may be checked by comparison of the surface and volume

Jan 1, 1965

By Richard L. Stein

INTRODUCTION Increasing a miner's chance of surviving a disaster requires advance planning for emergencies, adequate training of personnel, and the provision of proper survival equipment. It is extremely important that fires and toxic gases be detected early, and that warning systems such as stench, audible and visual alarms, and communications systems be provided to allow timely escape. However, situations do occur where timely warning is not provided and immediate escape is not possible, regardless of the precautions that are taken. Under such circumstances, a combination of new per¬sonal protective devices and old techniques can provide protection that is adequate to save a miner's life. SELF-RESCUE EQUIPMENT Self-rescuers are devices which provide a short-term supply of respirable air under the potentially lethal con¬ditions following a mine disaster. Functioning either as filtration devices for the elimination of carbon monoxide (CO) or as oxygen supplies, self-rescuers allow an endangered miner a limited amount of time in which to effect an escape or reach a place of temporary safety. Carbon-Monoxide Self-Rescuers For almost 50 years, filtration self-rescuers have protected miners' lives following explosions or fires. Designed to eliminate carbon monoxide from inhaled air, the first of these devices was approved in the 1920s and has evolved into the two units that are approved today-the Mine Safety Appliance (MSA) W-65 and the Draeger 810. Both of these units rely on the cata¬lytic conversion of very toxic carbon monoxide to rela¬tively safe levels of carbon dioxide (CO,). As shown in Figs. 1 and 2, both of these filtration self-rescuers operate in the same manner. On inhalation, the mine air passes through both coarse and fine filters that remove the dust, preventing the dust from coating the chemical beds or entering the miner's mouth. The air then passes over a drying agent that removes water vapor; this is required to prevent the water vapor from poisoning the catalyst. Subsequently, the dried air passes over the Hopcalite catalyst, where carbon monoxide is converted to carbon dioxide. Then, the air passes through a heat exchanger that reduces the temperature of the inhaled air. Finally, the air is inhaled by the user. Air exhaled by the user passes through the heat ex¬changer and exits the device through an expiratory valve. Both units of this type are subject to certain limita¬tions: 1) The units do not protect the user against oxygen¬deficient air; when air is inhaled through this type of self-rescuer, the device only removes the carbon mon¬oxide. If the mine air contains less than 15% oxygen, anoxia is inevitable. Symptoms of anoxia include dizzi¬ness, shortness of breath, quickened pulse, and deeper and more rapid respiration while the victim is at rest. During heavy exertion such as would be expected dur¬ ing escape efforts, a 15% oxygen level can cause loss of consciousness. 2) The units do not protect against excessive levels of carbon dioxide. Available data on carbon monoxide and carbon dioxide concentrations following explosions or fires are limited, but there are indications that the carbon monoxide concentration can rise to 2% and the carbon dioxide concentration can reach 5 or 6% of the mine air (by volume). In some situations, the in¬haled air could contain up to 6 or 7% carbon dioxide. At a carbon dioxide concentration higher than 2%, breathing patterns can be affected adversely. At a con¬centration of 6 or 7%, the effects include severe respira¬tory distress, with unconsciousness resulting from exer¬tion such as that needed for escape activities. 3) The units do not protect against high inhalation temperatures. The catalytic oxidation of carbon mon¬oxide to carbon dioxide is an exothermic reaction that evolves a large amount of heat. As the carbon monoxide concentration increases, the temperature of the inhaled air also increases. Available data show that above 1.5% carbon monoxide, the air inhaled through some self¬rescuers can be as high as 90°C (194°F). At those

Jan 1, 1982

By L. W. Legerwood

During the past three decades, the oil industry has expended increasing eflorts seeking improved drilling tools or systems to reduce drilling costs. The total cost of these efforts is unknown, but it certainly amounts 10 tens of millions of dollars. Most of the "new" system that past and present investigators have sought to develop actually are old public information. In seeking to implement new-systern concepts, investigators have tested the following: impact at frequencies ranging from 6 to 300 cycles per second; electrical, mechanical and hydraulic meam of actuating percussors; bit rotary speeds up to 2,000 rpm; electric and hydraulic bottom-hole means of rotating bits; bottom-hole machines with power outputs up to 400 hp; shock waves; explosives; high-velocity pellets; flame; arc; grinding wheels; abrasive jets; erosion by high-velocity gases; chemical attack; electric current; magnetic waves; retractable rock bits; reelable drill pipe; continuous coring with reverse circulation; and automation of drilling rigs. Table I shows how these investigations are grouped for discussion purposes in this paper. In spite of these efforts to discover new and improved systems, rotary drilling main;aim its economic leadership. Undoubtedly, rotary drilling costs will continue to be reduced by rigid application of the best available technology and by development of new rotary technology. In view of the extensive past development programs, however, significant long-range improvement appears to be a research, not a development problem. Research must postulate and prove theories and principles governing variour subsurface rock-failure processes pertinent to both rotary and new systems. Ah, research must produce physical and engineering data relative to these processes. When such information is available, earth boring will graduate from an art to a science. Major improvements in rotary drilling can then be expected, and the systematic evolution of an improved drilling method can be initiated—with a strong probability for success. Development of drilling methods other than rotary drilling has been one approach investigated by the industry as a means for reducing drilling costs. Major cost-reduction efforts, however, have been centered on engineering development work aimed at incremental improvements in rotary drilling. A relatively minor effort has been expended to establish basic physical principles and engineering data pertaining to the earth-boring process which can serve as a foundation for the development of cost-cutting drilling hardware. Current oil-industry economic trends have added impetus to the need for effective programs to reduce drilling costs. However, areas for expanded future efforts should be selected only after careful study of past investigations. It is the purpose of this paper to review past and Dresent industry efforts to develop new drilling tools or systems and to suggest areas where additional science is needed. Past reviews of drilling methods have been limited to a few processes prominent at the time the reviews were made. This paper seeks to present a concise, well organized review of all methods that have received actual development or test work. The preparation of a paper seeking to review all past developments is fraught with many difficulties, not the least of which are errors in or obsolesence of printed matter used as source material and the lack of data on industry developments which have not been published. Since de-

By Richard L. Bullock, Bruce Kennedy

Vince Lombardi once said, "Practice doesn't make perfect, perfect practice makes perfect." When it comes to building a mine that will operate at the optimum level for the set of geologic conditions from which it was developed, Lombardi's remark might be paraphrased to describe the problem: planning won't guarantee the best possible mine operation unless it is the best possible mine planning. Any sacrifice in the best possible mine planning introduces the risk that the end results may not reach the optimum mine operation desired. This section addresses many of the factors to be considered in the initial phase of mine planning. These factors have the determining influence on the mining method, the size of the operation, the size of the mine openings, the mine productivity, the mine cost, and, eventually, the economic parameters used to determine whether or not the mineral reserve even should be developed. A little-known fact, even within the metal-mining community, is that room-and-pillar mining accounts for most of the underground mining in the united States. According to a 1973 study on noncoal mining (Anon., 1974), more than 76% of the producing mines [of over 1089 t/d (1200 stpd) capacity] produced approximately 70 000 000 t (77,000,000 st) or 60% of the nation's underground tonnage of material by room-and-pillar mining. That same year, 96.8% of the nation's under- ground coal mines produced 262 950 000 t (289,911,000 st) of coal extracted from room-and-pillar mines (Anon., 1976). Thus, nearly 333 000 000 t (367,000,000 st) of the United States' raw material is produced from mines using some form of the room-and-pillar mining system. Because approximately 90% of all mining in the United States is done by some variation of room-and- pillar mining, it is appropriate to give special emphasis to the effects of the various elements of mine planning on room-and-pillar mining. The relationship of these elements to other mining methods will become apparent as the elements are described in later sections herein. TECHNICAL INFORMATION NEEDED FOR PRELIMINARY MINE PLANNING Assuming that the reserve to be mined has been delineated with diamond-drill holes, the items listed in the following paragraphs need to be established with respect to mine planning for the mineralized material. Geologic and Mineralogic Information The geologic and mineralogic information needed includes the following: 1) The size (length, width, and thickness) of the areas to be mined within the overall area to be considered, including multiple areas, zones, or seams. 2) The dip or plunge of each mineralized zone, area, or seam, noting the maximum depth to be mined. 3) The continuity or discontinuity within each of the mineralized zones. 4) Any swelling or narrowing of each mineralized zone. 5) The sharpness between the grades of mineralized zones within the material considered economically minable. 6) The sharpness between the ore and waste cutoff, including whether this cutoff can be determined by observation or must be determined by assay or some special tool; whether this cutoff also serves as a natural parting resulting in little or no dilution, or whether the break between ore and waste must be induced entirely by the mining method; and whether or not the mineralized zone beyond (above or below) the existing cutoff represents submarginal economic value that may be- come economical at a later time. *7) The distribution of various valuable minerals making up each of the minable areas. 8) The distribution of the various deleterious minerals that may be harmful in processing the valuable mineral. 9) Whether or not the identified valuable minerals are interlocked with other fine-grained mineral or waste material. 10) The presence of alteration zones in both the mineralized and the waste zones. Structural Information (Physical and Chemical) The needed structural information includes the following: * 1 ) The depth of cover. 2) A detailed description of the cover including: the type of cover; * the structural features in relation to the mineralized zone; * the structural features in relation to the proposed mine development; and * the presence of and information about water, gas, or oil that may be encountered. 3) The structure of the host rock (back, floor, hanging wall, footwall, etc.), including: * the type of rock; * the approximate strength or range of strengths; * any noted weakening structures; * any noted zones of inherent high stress; noted zones of alteration; the porosity and permeability; * the presence of any swelling- clay or shale interbedding; the rock quality designation (RQD) throughout the various zones in and around all of the mineralized area to be mined out; the temperature of the zones proposed for mining; and the acid generating nature of the host rock. 4) The structure of the mineralized material, including all of the factors in item 3 plus: * the tendency of the mineral to change character after being broken, i.e., oxidizing, degenerating to all fines, recompacting into a solid mass, becoming fluid, etc.; * the siliceous content of the ore; the fibrous content of the ore; and the acid generating nature of the ore. Economic Information The needed economic information includes: *1) The tons of the mineral reserve at various

Jan 1, 1982

By R. W. Carpenter, J. C. Ogle

It is possible, by using pack-rolling instead of conventional rolling, to reduce a number of metals to thicknesses of 2µm or less. Such thinfoils are generally made at room temperature without intermediate annealing. In addition, pack-rolled foils fail by developing pinholes at thicknesses near 2µm instead of developing the shear cracks usually observed in cold-rolled ductile metals. This paper presents the results of a general investigation of the deformation substructure and texture developed in copper and iron pack -rolled from 130 to about 2µm thickness. Electron microscopy showed that in both metals a fine (0.2 to 0.5?µ m) deformation subgrain structure formed during pack-rolling; in neither case was this substructure grossly different from substructures formed during conventional rolling. The deformation texture formed in pack-rolled iron was quite similar to usual bcc textures; however, in the case of copper, the cube texture was stable during pack-rolling and the normal copper deformation texture was unstable. It is shown analytically that the constraining pack induced a large hydrostatic pressure in the foils during pack-rolling. The pinhole failure mechanism is attributed to the presence of the large hydrostatic pressure during pack-rolling; this strongly suppressed the growth of shear cracks. The stability of the cube texture in copper is also probably due to the unusuul stress distribution developed during pack-rolling. EXPERIMENTS at several laboratories have shown that very thin foils of the common structural metals and many of the rare earths can be made by "pack-rolling". 1-3 The technique was originally developed to make specimens for nuclear scattering experiments and foils for X-ray filters. It is also useful for making experimental laminar metallic composite bodies and foils thin enough for direct examination by ultra-high voltage electron microscopy without the need for special thinning techniques. Pack-rolling in the present context means a three-layer pack, with the material to be rolled into foil comprising the center layer. The outer two layers, which constrain the foil during reduction, are ordinarily austenitic stainless steel. Typically, a 130 µm (0.005 in.) metal strip can be reduced to a final thickness of 2 µm or less by this process. This is accomplished at room temperature, without intermediate annealing. It has been observed that foils produced by this process do not exhibit at any stage of their reduction the severe work-hardening found in strip rolled by conventional cold-rolling methods. Neither is the failure characteristic the same."&apos; Conventionally cold-rolled ductile metal strip fails by developing shear cracks on planes whose normals nearly bisect the angle between the rolling direction and normal to the rolling plane; these are planes of maximum shear stress. In pack-rolling this mechanism has not been observed; failure occurs by the formation of pinholes on the foil surface (penetrating the foil). If pack-rolling is continued the hole density increases. These differences in behavior imply the existence of appreciably different substructure in pack-rolled foils compared to substructure in conventionally rolled material, or perhaps that the geometry of pack-rolling has an effect on the foil behavior. This paper describes an investigation of deformation substructure and texture in some specimens of pack-rolled copper and iron, and some considerations of the stress distribution in the foils during rolling that result from the geometry of pack-rolling. EXPERIMENTAL DETAILS Three different materials were used for pack-rolling in the present work: soft copper sheet (99.8 pct Cu, 0.03 pct 0, electrolytic tough pitch) and two types of iron, Ferrovac E* and Armco iron. Each was "Crucible Stccl Co. initially in the form of 130 µm annealed strip with grain size ranges of approximately 10 to 40 µm. The initial texture of the copper (determined as noted below) was the normally observed cube type (001)[100]; there was evidence of a small amount of material in the cube-twin orientation reported by Beck and Hu.4 The initial texture of the Ferrovac E was similar to that reported for recrystallized iron by Kurdjumov and sachs,5 who list the principal orientations as {111}<112>, {001}<110> 15degfrom RD and a weak component {112}(110) 15 deg from RD. The starting texture of the Armco iron was not determined. Pack-Rolling Procedure. A four-high mill was used for all specimens. The work roll and backing roll diameters were 1.625 and 5.25 in., respectively. The peripheral roll speed of the work rolls was about 2.5 in. per sec. All foils were initially reduced from 130 to 100 µm by conventional straight rolling and then inserted into a pack, without any intermediate annealing, for further reduction. The pack consisted of an 0.033 in. (838 µm) thick 3 by 6 in. polished sheet of austenitic stainless steel, folded to make a 3 by 3 in. jacket. After folding, the jacket was given a small reduction to close the fold tightly before insertion of the foil. During pack-rolling a constant change in roll spacing was made every third pass. The roll-spacing change corresponded to a 5 pct reduction in thickness for a new pack. This approached a 10 pct reduction when the pack had decreased to about half its original thickness. At this point the deformed pack was discarded and a new one

Jan 1, 1970

By David J. Mack, Dennis R. O’Boyle

The mechanical properties of three polycrystalline intermediale Phases that have the y bvass structure were measured in compression between 400° and 900°K. At the lower testing temperatures— termed Region I— no plaslic deformalion occurred prior to brittle fracture. A1 higher temperatuves— lermed Region III—the three phases deformted plaslically for all strain rules from 1.5 to 36.0 x 10-4 sec-1. The homologous temperature for the inilialion of plaslic flow increased in the same order as tile atomic size difference between atoms in each plmse (0.50Tmp for y AgZn, 0.62Tmp for y CuZn, and 0.76 Tmp for y CuCd). The increased flow resistance oj. y CuCd above 0.50Tmp is explaitzed in terms of the more random arrangement of copper and cadmiwn atoms on the laltice sites of&apos; the y bvass slrilctlire. Plaslic deformation of the three intevmiediate phases in the smooth plaslic flow region appears to be controlled by oacancy-induced dislocation climb. RESULTS are reported in this paper on the mechanical properties of three binary intermediate phases having the complex cubic y brass structure (D82, 143m) containing fifty-two atoms in the unit cell. The structure of the prototype y brass phase, y CuZn, was first analyzed by Bradley and Thewlis&apos; based on the X-ray work of Westgren and Phragmen.2 The y brass structure can be visualized as an arrangement of twenty-seven body centered cubes (3 by 3 by 3) into a unit cell containing fifty-four atoms. From this cell the center atom and the four corner atoms are removed and the remaining fifty-two atoms are slightly displaced, resulting in the y brass structure. Approximately thirty-five binary intermediate phases that have an electron to atom ratio of 21:13 have been reported to form the y brass structure. The three y brass intermediate phases selected for this study are formed from elements in groups IB and IIB of the first and second long period. Each compound exists over a composition range that includes the stoi-chiometric composition A5BB (Cu5Zn8, Ag5Zn8, Cu5Cd8). y CuZn (a0 = 9.944A) has solubility limits extending from 57 to 67 at. pct Zn at 500°C apd melts peritecti-cally at 834°C; y AgZn (a0 = 9.326A) has a solubility range from 58 to 63 at. pct Za and melts peritectically at 661°C. y CuCd (a, = 9.596A) melts congruently at 563°C and has a maximum range of solubility extending from 58 to 64 at. pct Cd. In order to minimize coring and microporosity during solidification, the composi- tion of the compounds that form peritectically (y CuZn and y AgZn) was chosen so that the temperature difference between the liquidus and the solidus is only a few degrees centigrade. In selecting y brass compounds for this study, phases were avoided that contained a transition element or that had an electron to atom ratio greater than 1.70. Hume-Rothery et a1.3 has observed that, when the electron to atom ratio of y brass intermediate phases exceeds 1.70, atoms begin to drop out of the unit cell to maintain a constant electron to unit cell ratio. EXPERIMENTAL PROCEDURE Each of the three intermediate phases was prepared from elements having a purity >99.999 pct. Only trace quantities of impurities (<0.0001 pct) were detected by spectrographic analysis of the four elements. A master alloy of each compound weighing approximately 130 g was prepared by melting the elements in evacuated sealed quartz tubes heated in an electric furnace. Specimens having the desired diameter for the mechanical property measurements were obtained by remelting the master alloy in a Vycor tube under argon and drawing specimens from the melt into 4-mm-diam quartz tubes. To obtain a smooth surface on specimens drawn from the melt, the inside of the quartz tube was coated with a thin layer of graphite formed by thermally decompositing acetone. All specimens were imbedded in sealing wax and cut to the proper length with a diamond-impregnated cut-off wheel. The individual specimens (4 mm diam by 8 mm long) were then sealed in evacuated Pyrex tubes and homogenized at 500°C for 72 hr. Chemical composition, variations over the length of the rod (10 cm long) were generally less than 0.2 pct and the analyzed composition differed from the desired composition by less than 0.5 pct.4 Metallographic evaluation of the specimens after homogenization showed no evidence of a second phase in any of the compounds. The average grain size of the specimens after homogenization was 1 to 2 mm. All mechanical property measurements were carried out in an argon atmosphere using a compression apparatus described previously.5 Prior to applying the load to the specimen, the test apparatus was evacuated, back-filled with argon, and heated to the test temperature for 20 min to establish thermal equilibrium. EXPERIMENTAL RESULTS Mechanical properties measured as a function of temperature and strain rate for each of the three y brass phases were similar. From room temperature to the ductile-to-brittle transition temperature, only elastic deformation was observed prior to fracture.

Jan 1, 1968

By W. C. Hellyer, R. A. Campbell

In single-stage autogenous grinding, the buildup of a critical size fraction in the media can be corrected by removing this material through pebble ports, crushing it below the critical size range, and recycling the crusher product back to the mill. The rate at which the critical sire fraction is crushed affects the size distribution of the grinding media and this in turn affects the sire distribution of the grate discharge. This provides a means for controlling the grind produced by a single-stage autogenous grinding unit. The pilot-plant investigations on a very hard copper ore were carried out at the facilities of the Institute of Mineral Research, Michigan Technological University. In an autogenous grinding circuit in which feed at approximately 9 in. top size is reduced to a size suitable for subsequent processing, the build-up of a "critical size" fraction in the media causes problems. A "critical size" fraction has been defined as, "media too small to effect reduction by impact grinding of ore coarser than a quarter of an inch and too large to be broken by the largest size of media in the charge."&apos; The buildup of a critical size fraction reduces the capacity of the mill, increases the grinding power requirements per ton of finished product, and generally produces a finer grind than is desired. It is general practice to overcome this problem by the addition of large-diameter steel balls to the grinding charge. This has certain disadvantages, such as 1) an increase in mill liner wear, 2) wear on the steel balls, and 3) some loss in flexibility in grinding circuit opera-tions resulting from difficulties in removing the steel balls by means other than by grinding out. A number of investigators&apos; have suggested the use of a small crusher as an alternative to the use of large steel balls for control of a critical size fraction. With this technique the critical size fraction is removed from the mill continuously through suitable-sized pebble ports, crushed below the critical size range, and returned to the mill. An external crusher would cause no increase in mill liner wear, the wear on the crusher would be less expensive than the wear on the steel balls, and the flexibility of the grinding circuit operations would be enhanced since the crusher can be cut into and out of the circuit at will. This paper describes an autogenous grinding pilot plant investigation on a very hard copper ore, which led to the selection of an autogenous grinding flowsheet incorporating a crusher in the circuit as the preferred method for grinding this ore. Further development of the technique demonstrated that the crusher could be employed to control the product produced by an autogenous grinding unit. Description of Pilot Plant These investigations were carried out at the pilot-plant facilities of the Institute of Mineral Research, Michigan Technological University, Houghton, Mich. The autogenous grinding unit was a 6-ft-diam by 2-ft-long Hardinge Cascade mill with %-in. slotted steel grates. Pebble ports cut into the grates allowed passage of pebbles having a top size of approximately 21/2 in. The grate and pebble-port discharge material passed over a 3/16-in, trommel screen with the trommel under -size being pumped to the classification device. DSM screens were employed for classification in the early investigations, but were replaced later by a small Dorr rake classifier. The trommel oversize material returned to the mill via scissor conveyors. When the crusher was incorporated into the circuit, the trommel oversize material passed to a double-deck vibrating screen fitted with 2-in. and 3/4-in. square mesh screen cloths. The $ 2-in. and the —2 +3/4-in. size fractions were combined and crushed to a nominal 1 in. in a 4 x 6-in. jaw crusher. The crusher discharge and the —3/4-in. screen undersize returned to the mill. In the final investigation, a flap gate fitted to the top deck of the screen directed the — 21/2 +2-in. size fraction to the crusher or back to the grinding mill. The flap gate operated manually on a 15-min cycle, with this material being crushed for so many minutes out of each cycle. This was found to be a more reliable method of controlling the amount of — 21/2 +2-in. material crushed than attempting to make a split of a small weight of material on a continuous basis. Operational Techniques: Each sample was sized into three fractions and the Cascade mill feed was reconstituted from these size fractions in the proportions existing in the original sample. Sufficient feed for 15 min operation was weighed out and fed by hand over a 15-min period. The gross mill power draft was recorded every 15 min, corrected for tare power and drive efficiency, and reported as net kilowatt-hours per ton. A recording kilo-watt-hour meter provided a continuous visual record of the power drawn by the mill. Pulp densities of the trommel undersize and the classifier overflow (or DSM undersize) were taken every 15 min. The classifier overflow was sampled automatically. Timed samples of the classifier sands were taken every 15 or 30 min, weighed, and returned to the circuit. After correcting for moisture content, the weights were converted into percent circulating load. Timed samples of the +2 in., the —2 + 3/4-in., —3/4-in. screen fractions, and the crusher discharge, were

Jan 1, 1971

By D. J. Lachel, Douglas E. Hansen, Bruce Kennedy

The very high cost of exploration and development for new mines requires that a carefully planned, complete, and technically competent geotechnical evaluation be applied to the ore body and the surrounding rock. The ground conditions must be predetermined sufficiently so that, during the development and exploitation stages, unexpected conditions will be of minor consequence. Ground conditions are a major factor affecting the cost of excavated items such as shafts and inclines. For example, shafts cost less to sink in tight dry ground than in highly pervious ground below the water table, with all other factors being equal. The proper mining methods also is dependent upon the ground conditions. Bad ground for room-and-pillar mining may be ideal for caving methods; whereas, ideal ground for room-and-pillar mining would be extremely difficult for a caving operation. To do a proper engineering design, the parameters of strength, deformability, and permeability must be determined, in addition to the loads that will be applied as a result of the mining activity. The required degree of accuracy for each parameter depends upon the individual case. For example, the strength of massive rock in which openings of small cross-sectional area are to be driven at shallow depths needs to be resolved only within several thousand kilopascals (pounds per square inch). Under such conditions, the rock strength is not likely to be of great economic importance unless the rock is very weak. Large openings at great depths require a much more accurate appraisal of rock strength and deformation characteristics. Ground water conditions are important at any depth. Today, a number of techniques exist for evaluating conditions before opening up the ground for development. These techniques and their applications are described in more detail later in this text. The evaluation of ore body ground conditions should start at an early stage, as soon as the deposit has been proven viable for a mine. Economic geology investigations mainly serve to define the value and limits of the ore deposit. To do this properly, the geology of the site is determined in at least as much detail as is necessary to define the "controls" for the ore body. The ground-condition evaluation must determine the engineering properties of the ore body and the surrounding material. This includes the structural geology as it relates to rock strength; i.e., the frequency, extent, and orientation of discontinuities such as faults, joints, bedding, etc. Ground water conditions also must be determined. It is important that the detailed geology of the site be determined at an early date. Reasonable geotechnical parameters must be established during this time period, so they can be used in mining model studies that include both physical and mathematical models. The ultimate goal bf-the geotechnical evaluation is to predict the ground behavior as the excavations are made and how the ground behavior will affect the safety and economic well-being of the mine. The formulation of a ground-condition evaluation program should be planned carefully to determine the particular geotechnical parameters only in as great a detail as will have a bearing on the mine design. For instance, it is not necessary to make numerous triaxial tests in strong rock when the mining will be conducted within a few meters (feet) of the surface in a regime of gravity loading and an absence of tectonic stresses. Similarly, it is not necessary to invest large amounts of time and effort in the determination of joint strengths if the orientations of the joints are such that the failure could not occur along them. Generally, the determination of ore body ground conditions can be divided into four main steps: 1) The general and structural geology of the site must be determined; this is determined primarily during the economic evaluation. 2) The material behavior characteristics of the deposit and the vicinity must be determined. 3) The ground water hydrology of the deposit and the vicinity must be determined. 4) Miscellaneous ground support considerations must be established. Gross features are defined first, followed by a series of increasingly detailed investigations. Specifically, the geology study should start by utilizing techniques such as remote sensing, topographic map interpretation, geologic mapping, etc. After the gross features, such as major faults, have been delineated, the studies are directed toward defining the more detailed elements of jointing, rock strength, ground water hydrology, and other characteristics. This information is used to formulate an assessment of the strength of the ore body and the surrounding rock. Finally, all of the gathered information is integrated to make an assessment of the rock behavior in relation to the types of openings, the excavation methods, the support requirements, etc., that should be used for the rock conditions that are present. The mine design and the geotechnical assessment should be an iterative process. General conclusions about the rock conditions will dictate a type of mine plan. The specific mine plan then should be modified by specific geotechnical considerations. For example, an initial assessment may conclude that the ore body is strong and that a strong rock unit is present over the ore body. Thus, one definite possibility is to mine the deposit by room-and-pillar methods. However, subsequent investigations may reveal faults through the deposit, requiring relocation of the mine openings so that pillars and other structures are in sound rock. In other words, the final mine plan should be compatible with the geotechnical aspects of the ore body.

Jan 1, 1982

By Ursula E. Wolff

The precipitation of carbides from an alloy containing 33 pct Ni, 21 pct Cr, balance iron, was investigated electron microscopically by means of extraction replicas and thinned metal foils. Annealing temperatures ranged from 565°to 870°C and up to several thousand hours. M23C6 precipitated in pain boundaries, incoherent and coherent twin boundaries in that sequence. The orientation relationship between carbides and austenite matrix was determined and correlated with the morphology of the carbides and with the type of boundary in which precipitation occurred. In large-angle grain boundaries, as well as in coherent twin boundaries, the carbides had the same orientation as one of the adjacent pains. These carbides formed sheets of individual flakes with shapes related to the orientation of the boundary. In incoherent twin boundaries carbides precipitated in ribbons composed of pavallel rods. An unidentified subcarbide was found to precede precipitation of M23C6 in these boundaries. The M 23 C6 rods had a kind of fiber texture with (110) parallel to the long dimension of the rods and ribbon, and with orientations of both of the adjacent twin-related austenite crystals Predominant in the texture of the carbide. A hard sphere crystal model has been used to discuss orientation and morphology of the carbides in terms of free volume and vacancies available in the boundaries. A number of papers have dealt with the morphology of chromium carbide (M23 C6) precipitated in austenitic stainless steels.1"7 In all these investigations, the carbides were examined in the electron microscope by means of extraction replicas. With this technique, the carbides retain the spatial distribution they had in the bulk sample. However, since the matrix is dissolved in the process, the particles can turn in an unpredictable way; and the orientation relationship between matrix and carbides cannot be established. In this paper the results of studies on extraction replicas and on thinned metal foils are reported. These studies were undertaken to determine the matrix-to-car bide orientation relationship, and to correlate the orientation of the carbides with their morphology. PROCEDURE The material used was an austenitic alloy with 33 pct Ni, 21 pct Cr, balance iron, containing approximately 0.05 pct C. Coupons of 1.25-mm sheet were first solution-annealed at 1050°C for 15 min and air-cooled. Then, to precipitate the carbides, samples were isothermally annealed in the range from 565" to 870°C for times up to several thousand hours. All further specimen-preparation procedures were carried out after the final anneal. Carbon extraction replicas from polished and etched surfaces were made with 10 pct bromine in methyl alcohol.&apos; Thin foils were prepared from punched-out 3-mm-diam disksg which fit into the electron-microscope holder. The disks were prethinned by grinding to approximately 0.5 mm thickness, and then electro-polished in a polytetrafluoroethylene holder1&apos; with a solution containing 5 pct perchloric acid in acetic acid to which 10 g per 1 Cro3 and 5 g per 1 nickel chloride were added (etchant modified from that of Briers et al."). This solution dissolves neither the carbides nor the austenite around the carbides preferentially. By using extraction replicas, electron micrographs and selected-area electron-diffraction patterns were taken from the same carbide arrays. By using thin foils, electron micrographs were made from a grain boundary area containing carbides. Electron-diffraction patterns were then taken from the same area and from each of the adjacent grains separately. In this manner, the orientation of each grain could be determined without interference by the carbide pattern. A peculiarity of extraction replicas should be pointed out. After the matrix is etched away, the carbide arrays float freely in the etching and washing solutions, and are held in place only at the anchoring points in the carbon replica. When the replica is picked up with a screen the carbide arrays tend to flip to one side. Thus, while the surface features are preserved, the original arrangement of the carbides may severely and unpredictably be disturbed whenever the specimen contains large amounts of interconnected carbides. Nevertheless, it is possible to correlate the different morphologies of the carbides with the type of boundary in which they have precipitated. RESULTS 1) Extraction Replicas. Fig. 1 shows that the grain boundaries usually are curved, multicornered surfaces of random orientation. The coherent twin boundaries (which are (111) planes) cut a grain into parallel slices. Incoherent twin boundaries occur at the ends and on the steps of twins and are often narrow, parallel-sided strips which are much longer than they are wide. Different morphologies can clearly be distinguished for the M23Ce carbides precipitated in each of these types of boundaries, and agree well with those observed by kinzel.2 The kinetics of this precipitation has been investigated." The first carbides precipitate in junctions of three grain boundaries and fan out from there into the adjoining boundary surfaces, Fig. 2(a). These carbides are oriented randomly, Fig. 2(b), and become coarser and thicker as annealing time increases. The large-angle grain boundaries are next to fill

Jan 1, 1967

By O. D. Gonzalez

Existing measurements for the steady-state permeation of hydrogen in a iron above 100°C have been examined for contribution of determinate errors. The analysis leads to a recommended equation for the permeability of hydrogen in a iron: o= (2.9 ±0.5) x 10-3 exp - (8400 ± 400)/RT cu cm (ntp H2) cm-1 sec-1 atm-1/2 THE permeability of a iron to hydrogen has been the subject of numerous investigations over the past 40 years, and at present there are thirteen sets of published results for the rate of steady-state permeation of hydrogen in a iron above 100°C. The numerical values in each set of results are entirely self-consis-tent, but the spread among the sets is too large to be attributed solely to experimental error, i.e., to error other than in the specimen itself. Several reasons have been advanced to explain the disparities, but to date the relative importance of experimental inaccuracy to the spread remains uncertain. The purpose of this report is to examine in detail the sources of determinate errors inherent in the experiments and to assess as far as possible the contribution of the errors to the results. The ultimate goal is the selection of values for the permeability and heat of permeation most nearly representative of hydrogen in a iron. The analysis is limited to those experiments in which the permeation rate was observed at steady state—a condition in which traps for hydrogen within the metal are filled to a fixed level15 so that the trapping mechanism is not reflected in the rate of passage of the gas. Furthermore only data are examined in which surface processes are judged to have little or no influence on the flow. It is hoped with these restrictions to obtain values of the permeability and the heat of permeation which will be as closely related as possible to the mechanism of lattice diffusion. I) DEFINITION OF TERMS; UNITS In this report the data for permeation are given in terms of a coefficient oj permeability, ?, which is defined by the equation: jt=?A/?x{p1/2-po1/2) [1] where jt is the total flow of gas normal to the surface of a membrane of planar geometry, e.g., a disc, of area A and thickness ?x; pi and po are the pressures in the input and output sides, respectively. For flow radial to the walls of a membrane of cylindrical geometry, e.g., a tube, the corresponding equation is: where 1 is the length of the cylinder, and ri and ro are the inner and outer radii, respectively. The flux normal to the surface is given by Fick&apos;s law: j= -D(dc/dx) [3] At steady state the concentration gradient will be constant, and integration of Eq. [3] gives for the total flow through a disc of area A and thickness Ax: h =-DA(co - ci) [4] where c, and ci are the concentrations of solute at the output and input surfaces, respectively. When surface control is absent, co and ci are given by Sievert&apos;s law c = Kp1/2, and substitution therewith into Eq. [4] gives directly Eq. [I] where ? = DK. Integration of Fick&apos;s Eq. [3] in cylindrical coordinates will give Eq. [2] where again ? = DK and is thus shown to be independent of geometry (provided that surface control is negligible). The coefficient of permeability, or simply the permeability,* must be expressed in proper units. In *The term permeability will refer in this report always to the coefficient defined above; permeation will be used to specify the general phenomenon of gas passage through a membrane. this report ? will be expressed in the units of cu cm (ntp H2) cm-1 sec-1 atm-1/2. The variations of D and K with temperature are given by D = Do exp(-Ea/RT) and K = KO exp(-?Hs/RT) where E, is the activation energy for diffusion and AH, the heat of solution, each usually expressed in calories per mole of solute. The variation of permeability with temperature will thus be given (for conditions where surface control is negligible) by ? = ?o exp(-?Hp/RT) where ?0 = DoKo and ?Hp = Ea + ?Hs. The units of ?0 are the same as those of 6, and??Hp will be expressed in calories per mole H. 11) SUMMARY OF PERMEABILITY RESULTS Table I gives the values reported to date for the permeability of H2 in a iron in terms of ?o and ?Hp. Except where noted the parameters listed were taken directly from the numbers reported by the various investigators with only a change in units. The temperature limits within which the listed ?o and ?Hp hold are given in column 7; the limits marked in parentheses in this column indicate the entire temperature range covered in each investigation. The listed values of ?o and ?Hp are those giving a linear plot of ln? against T-1 at the higher temperatures in each set of measurements, and thus presumably represent the case for which surface control was negligible. Column 6 gives values of 9 at a representative tem-

Jan 1, 1970

By N. L. Svensson

The model developed by Tuylor for the calculation of Polycrystalline yield strength has been applied to the case of an aggregate hawing a preferred orientation. In general this procedure requires the specification of texture by means of weighting factors applied to specific orientations. The problem to which the model has been applied is that of the yield-strength aniso-tropy of cold-rolled aluminum whose rolling texture was described as a combination of (110)[112] and (311) [112] In this case yield-strength anisotropy is defined by the rutio of yield strength measured at an angle 8 to the rolling direction to that measured along the rolling direction. The method of calculation of yield-strength ratio as a function of ? is described and the results show good agreement with experimental values. The orthotropic yield criterion suggested by Hill has been applied to the results and the strain ratio R also calculated as a function of ?. This has been compared with calculations using the method suggested by Elias, Heyer, and Smith which does not exhibit suck good agreement with observation. one deficietlcy of the method presented is that the strain ratios used by are those applying to iso-Irobic materials. The method should therefore be reg-clrded only as a first abbroximation to the prediction of anisotropy. THE problem of calculating the stress-strain characteristics of polycrystalline aggregates from the properties of single crystals has attracted attention for a number of years. The most important contributions to this study have been those due to: Sachs,&apos; Cox and sopwith,2 Taylor,3 Kochendorfer,4 Batdorf and Budiansky,5 Calnan and Clews,6 Bishop and Hill,7,8 Kocks,9 Budiansky, Hashin, and sanders, 10 Kroner,11 Cyzak, Bow, and payne, 12 Budiansky and Wu,13 and Lin.14 While the earlier work has been largely superseded, recent developments tend to support Taylor&apos;s solution" within the restriction imposed by his assumptions. The essential features of Taylor&apos;s approach were: 1) the material is rigid-plastic; 2) each grain experiences the same strain components as the aggregate as a whole (the problem was that of uniaxial deformation with principal strain components in the ratio 3) all regions of each grain deform uniformly; 4) work hardening occurs equally on all slip systems. While Bishop and Hill7 have generally validated this approach, there has been some criticism offered. Kocks? as pointed out that since multiple slip must occur the single-crystal data must be determined from orientations arranged such that polyslip takes place. Boas and Hargreaves,15 and others, have shown experimentally that the strain distribution within grains is not uniform, the strains in the vicinity of grain boundaries being less than those in the center of the grains. Both of these criticisms can be largely offset by the suitable choice of single-crystal critical shear stress. However, for the problem analyzed below, the critical shear stress is not directly used and, consequently, these criticisms lose their importance. The more recent contributions have attempted to obtain a more complete analysis by considering an elas-toplastic material and considering interactions between grains of differing orientations. Lin14 has considered the early stages of yielding for a polycrystalline aggregate having specific regions of defined slip plane orientations. On the other hand, Budiansky and Wu13 have allowed for these interactions for randomly disposed grain orientations and have calculated the polycrystalline stress-strain curves for crystals exhibiting either elastic-ideally plastic or kinematic hardening characteristics. This work has shown that yielding commences when the macroscopic stress is 2.2 times the critical shear stress for slip in a single crystal (7,). The yield stress-strain curve then rises becoming asymptotic to a value of 3.072 7,. This is close to the value obtained by Bishop and Hill (3.06) in their confirmation of Taylor&apos;s method. This, of course, is to be expected since, at large strain values, the elastic strains are negligible and the rigid-plastic model is satisfactory. The results of Budiansky and Wu indicate that the result obtained by Taylor is 7.7 pct high at a plastic strain which is two times the elastic strain at the initiation of yield. By defining the anisotropy in terms of relative values, the ratio of yield strength at orientation ?, to that measured in the rolling direction, the effect of the discrepancy in Taylor&apos;s solution is considered to be of lesser consequence. Therefore, it is anticipated that an analysis based on Taylor&apos;s solution, which can be quite straightforward, should provide a reasonable estimation of the anisotropy of materials having a preferred orientation texture. OUTLINE OF TAYLOR&apos;S METHOD In fee metals there are four possible slip planes (the octahedral planes) and in each there are three possible slip directions (the edges of the octahedron), that is a total of twelve possible slip systems. von Mises16 has shown that at least five independent slip systems must become operative in each grain of the polycrystalline aggregate in order to preserve continuity of strain. With this geometrical requirement as basis and the assumptions previously listed, Taylor determined the operative slip systems for a number of orientations of the tensile stress axis specified in the unit stereographic triangle. For the ith slip system, the critical shear stress

Jan 1, 1967

By J. Chipman, T. Fuwa

The effects of various elements on the activity coefficient of carbon in liquid iron have been studied by two experimental methods: 1) equilibration with controlled mixtures of CO and CO2; 2) the solubility of graphite in the melt. Activity coefficient of C is increased by Al, Co, Cu, Ni, P, Si, S, and Srz. It is decreased by Cr, Cb, Mn, Mo, W, and V. THE thermodynamic properties of the iron-carbon binary system have now been fairly well established, although some uncertainty remains with respect to the exact location of some of the phase boundaries. The activity of carbon in ferrite and in austenite has been measured in the classic researches of R. P. smith&apos; while similar measurements by Richardson and ~ennis, and by Rist and chipman3 have established the values of the activity of carbon in liquid iron up to 1760°C. On the other hand, our knowledge of the effects of alloying elements on the activity of carbon in dilute solutions is restricted to Smith&apos;s experiments on systems Fe-C-Mn and Fe-C-Si in the austenitic range and to some more recent experiments of schwarzman4 in the a range. In addition there have been a number of determinations of the effects of various elements on the solubility of graphite in liquid iron, and from these the corresponding effect in saturated solution may be obtained. The purpose of the present study was to extend the investigation of the liquid system to include the effects of alloying elements upon the activity coefficient of carbon, principally in dilute solutions. Equilibrium measurements were made on the reaction C + co, = 2 CO (g) The prepared mixture of CO and CO,, diluted with argon, flowed over the surface of the liquid metal which, after several hours&apos; exposure to the gas, was quenched and anqlyzed. As in the earlier experiments, the principal experimental difficulty was in the deposition of carbon on the parts of the furnace at temperatures slightly below that of the metal bath. In order to minimize this difficulty, the ratio (Pco)2 /PCo2 was restricted to values not much higher than 100 atm, and correspondingly the carbon concentration in the metal seldom exceeded 0.30 pct. EXPERIMENTAL METHODS The method and apparatus were essentially the same as used by Rist and Chipman.3 The gaseous mixture consisting of highly purified CO, CO,, and argon, each controlled by a flowmeter, was led into the furnace and passed over the surface of the liquid-iron melt which was heated and stirred by high-frequency induction. One slight modification was made in that a molybdenum susceptor was placed outside the crucible for the sake of uniformity of temperature and to combat the tendency of carbon to precipitate on the crucible wall. Pure alumina crucibles approximately 25 mm ID were used. The charge consisting of about 30 g was made up of electrolytic iron, the alloying element to be added, and enough graphite to supply slightly more or less than the anticipated equilibrium carbon concentration. All metals used were of high purity. Metallic chromium, columbium, and vanadium were from special lots supplied by the Electro Metallurgical Co. Tin, copper, molybdenum, tungsten, cobalt, and nickel were of purest commercial grades. The electrolytic iron, after being cut to the proper size for charging, was prereduced by hydrogen at 850° to 1000°C to remove surface oxidation. The oxygen content of the reduced material was 0.002 pct. This treatment made it easy to control the carbon content of the initial melt. The charge was melted under the gas mixture to be used for the entire run. In some earlier melts the charge was melted under a stream of argon, but in this case some alumina was reduced from the crucible, and the aluminum thus absorbed in the melt was subsequently oxidized with the formation of a solid film of alumina on the surface of the melt. AS another safeguard against film formation, overheating of the bath was carefully avoided. All runs were made at a temperature of 1560°C. Under experimental conditions a charge of pure iron picked up 0.17 pct C in 3 hr and 0.23 pct C in 6 hr under an atmosphere for which the equilibrium concentration of carbon is 0.27. It is clear that the time required to reach equilibrium from an initially carbon-free melt would be very great. For this reason each experiment was started with a melt of known carbon concentration not far above or below the expected equilibrium value, and each melt was held at temperature for a period of at least 5 hr. Under such circumstances it was possible to chart the approach to equilibrium from both high-carbon and low-carbon materials. Temperature was controlled by frequent optical observation and adjustment and the metls were timed in such a way that the final 2 hr occurred during a time when electric power was steady; for example, 2 to 4 pm or after 11 pm. In melts containine volatile metals such as copper, tin, and mangane\e the time of holding was decreased somewhat in

Jan 1, 1960

By H. S. Ten Eyck

Fertilizer normally is spoken of as having three main components: nitrogen, phosphorus and pot- ash. Certainly, however sulfur must also be considered a basic component of fertilizers, even though in many instances it does not end up as a part of the end product. In addition, with heavier cropping and the resultant drain of plant foods from the soil, we have now reached a point where the minor elements necessary for plant growth have become a major factor in the fertilizer industry.

Jan 7, 1967

By Roy Sage

Overview ? History ? background to current programs ? Kyoto status ? Support for technology development ? Support for GHG reduction ? Sources of information History ? 1988 - Toronto Conference - The Changing Atmosphere ? Proposed target ? CO2 levels reduced to 20% of 1988 levels by the year 2005 ? 1992 - Canada signed Framework Convention on Climate Change at the UN Conference on Environment and Development in Rio de Janeiro in June 1992 ? Canada launched the National Action Plan on Climate Change ? Kyoto protocol adopted in principle in 1997

Nov 1, 2002

By J. P. Sieverson

The Comprehensive Environmental Response, compensation, and Liability Act (CERCLA -commonly referred to as Superfund) was passed into law in December 1980 in the wake of such incidents as Love Canal and the Valley of the Drums. Superfund was originally perceived as a way of cleaning up spills of hazardous substances and/or cleaning up abandoned hazardous waste dumpsites. In February 1981, Michael Cook, EPA deputy assistant administrator for hazardous emergency response, called Superfund "a tremendously powerful law. It could revolutionize the way the environment is protected, but we certainly don&apos;t intend to realize the full potential of the law for some time." These words have proven prophetic for Asarco&apos;s East Helena Smelter.

Jan 1, 1988

By Peter N. Calder, Alan Bauer

The complexity of modem open pit load-haul-dump systems is described, and the need for realistic working models is pointed out. Both standard and probabilistic simulation procedures are discussed, probabilistic simulation being indicated as the more powerful and practical technique. Acquisition of the necessary pit operating data required for simulator construction and general modelling procedures are described. Finally, results from a typical open pit simulator are used to indicate the scope of the technique.

Jan 1, 2014

By B. Hearn, O&apos

Falconbridge, Ltd. operates six mines (the strathcona, Fraser, Lockerby, Onaping, East and Kidd Creek) in the Sudbury and Timmins districts of northern Ontario (Fig. 1). These mines vary in depth to approximately 1500m and experience a wide variety of ground conditions. Problems of greatest concern include rockbursting, backfill pillar stability, wall and pillar stability, and the use of cable anchors. A study was initiated approximately two years ago to identify how numerical models could aid in the examination and possible understanding or solution of these problems. This paper discusses the use of numerical models at Falconbridge and describes a case history of their use in examination of rockbursting at the Strathcona Mine.

Jan 1, 1988