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By J. D. Miller, J. E. Sepulveda

Tar sand deposits in the state of Utah contain more than 25 billion bbl of in-place bitumen. Although 30 times smaller than the well-known Athabasca tar sands, Utah tar sands do represent a significant domestic energy resource comparable to the national crude oil reserves (31.3 billion bbl). Based upon a detailed analysis of the physical and chemical properties of both the bitumen and the sand, a hot-water separation process for Utah tar sands is currently being developed in our laboratories at the University of Utah. This process involves intense agitation of the tar sand in a hot caustic solution and subsequent separation of the bitumen by a modified froth flotation technique. Experimental results with an Asphalt Ridge, Utah, tar sand sample indicated that percent solids and caustic concentration were the two most important variables controlling the performance of the digestion stage. These variables were identified by means of an experimental factorial design, in which coefficients of separation greater than 0.90 were realized. Although preliminary in nature, the experimental evidence' gathered in this investigation seems to indicate that a hot-water separation process for Utah tar sands would allow for the efficient utilization of this important energy resource. The projected increase in the ever-widening gap between the domestic energy demand and the domestic energy supply for the next few years has motivated renewed interest in energy sources other than petroleum, such as tar sands, oil shale and coal. Although a number of research programs on the exploitation of national coal and oil shale resources have already been completed, very few programs have been initiated on the processing of tar sand resources in the United States. In recognition of their significance as a domestic energy resource, investigators at the University of Utah have designed an extensive research program on Utah tar sands. An important phase of this program, and the main subject of this publication, is the development of a hot-water process for the recovery of bitumen from Utah tar sands, as a preliminary step toward the production of synthetic fuels and petrochemicals. The term "tar sand" refers to a consolidated mixture of bitumen (tar) and sand. The sand in tar sand is mostly a-quartz as determined from X-ray diffraction patterns. Alternate names for "tar sands" are "oil sands" and "bituminous sands." The latter is technically correct and in that sense provides an adequate description. Tar sand deposits occur throughout the world, often in the same geographical areas as petroleum deposits. Significantly large tar sand deposits have been identified and mapped in Canada, Venezuela and, the United States. By far, the largest deposit is the Athabasca tar sands in the Province of Alberta, Canada. According to the Alberta Energy Resources Conservation Board (AERCB),2,3 proved reserves of crude in-place bitumen in the Athabasca region amount to almost 900 billion bbl. To date, this is the only tar sand deposit in the world being mined and processed for the recovery of petroleum products. Great Canadian Oil Sands, Ltd. (GCOS) produces 20 million bbl of synthetic crude oil per year. Another plant being constructed by Syncrude Canada, Ltd. is expected to produce in excess of 40 million bbl of synthetic crude oil per year. According to the Utah Geological and Mineral Survey (UGMS), tar sand deposits in the state of Utah contain more than 25 billion bbl of bitumen in place, which represent almost 95% of the total mapped resources in the United States.4 The extent of Utah tar sand reserves seems small compared to the enormous potential of Canadian tar sands. Nevertheless, Utah tar sand reserves do represent a significant energy resource comparable to the United States crude oil proved reserves of 31.3 billion bbl in 1976.5 Tar sands in Utah occur in 51 deposits along the eastern side of the state.4 However, only six out of these 51 deposits are worthy of any practical consideration (Fig. 1). As indicated in Table 1, Tar Sand Triangle is the largest deposit in the state and contains about half of the total mapped resources. Information regarding the grade or bitumen content of Utah deposits is still very limited. The bitumen content varies significantly from deposit to deposit, as well as within a given deposit. In any event, the information available6-8 seems to indicate that Utah deposits are not as rich in bitumen as the vast Canadian deposits which average 12 to 13% by weight.9 Although many occurrences of bitumen saturation up to 17% by weight have been detected in the northeastern part of the state (Asphalt Ridge and P. R. Spring), the average for reserves in Utah may well be less than 10% by weight. Separation Technology As in any other mining problem, there are two basic approaches to the recovery of bitumen from tar sands. In one

Jan 1, 1979

By J. W. Spretnak, D. A. Chatfield, G. W. Powell, J. R. Eifert

In nzost cases, the driving force for a lattice transformation is produced by supercooling below the equilibriunz transformation temperature. The interfnce reaction in isothermally annealed, multiphase diffusion couples may involve a luttice transformation which also requires a driving force. Direct experinzental evidence has been obtained for the existence of the driring force in the form of a supersaturated phase at the aocc)-0@cc) interface in Cu:Cu-12.5 ult pct A1 couples; the super saturation is equivalent to an excess free energy of approximately 3 cal per mol at 905. A tentatiue interpretation of the dynanzic situation a1 the interface based on the free energy-composition diagram is proposed. THE presently accepted theory of diffusion in multiphase couples1 states that there will be a phase layer in the diffusion zone for every region which has three degrees of freedom and which is crossed by the diffusion path in the equilibrium phase diagram. For binary systems, this restriction excludes all but single-phase fields and, for ternary systems, only one- and two-phase fields are included. In addition, Rhines"~ as well as other investigators3 6 have reported that the compositions of the various phases adjacent to the interfaces are, for all practical purposes, the compositions given by the intersections of the diffusion path with the solubility limits of the single-phase fields of the equilibrium phase diagram. Some studies of the rate of thickening of these intermediate diffusion layers indicate that the thickness of the layer changes para-bolically with time, or: where x is the position of the interface relative to an origin xo, t is the diffusion time, and k is a temperature-dependent factor. crank7 shows mathematically that, if the compositions at an interface are independent of time and the motion of the interface is controlled by the diffusion of the elements to and from the interface, then the segments of the concentration penetration curve for a semi-infinite step-function couple will be described by an equation of the form: hence, Eq. [l] follows from Eq. (21 if the interface compositions are fixed and if the motion of the interface is diffusion-controlled. Although the concept of local equilibrium being attained at interfaces has assumed a prominent role in the theory of diffusion in multiphase couples, experimental evidence and theoretical discussions which challenge the general validity of this concept have been reported in the literature. arkeen' has stated that strict obedience to the conditions set by the equilibrium phase diagram cannot be expected in any system in which diffusion is occurring because diffusion takes place only in the presence of an activity gradient. Darken also noted that it is usually assumed that equilibrium is attained locally at the interface although the system as a whole is not at equilibrium, the implication being that the transformation at the interface is rapid in comparison with the rate of supply of the elements by diffusion. ISirkaldy3 indicates agreement with Darken in that he believes the concept of local equilibrium is at best an approximation because the motion of the phase boundary requires that there be a free-energy difference and, hence, a departure from the equilibrium composition at the interface. Seebold and Birks9 have stated that diffusion couples cannot be in true equilibrium, but the results obtained are often in good agreement with the phase diagram. The initial deviation from equilibrium in a diffusion couple will be quite large because alloys of significantly different compositions are usually joined together. Kirkaldy feels that the transition time for the attainment of constant interface compositions (essentially the equilibrium values) will be small, although in some cases finite. Castleman and sieglelo observed such transition times in multiphase A1-Ni couples, but at low annealing temperatures these times were quite long. Similarly, ~asing" found departures, which persisted for more than 20 hr, at phase interfaces in Au-Ni and Fe-Mo diffusion couples. Braun and Powell's12 measurements of the solubility limits of the intermediate phases in the Au-In system as determined by microprobe analysis of diffusion couples do not agree with the limits reported by Hiscocks and Hume-Rothery13 who used equilibrated samples. Finally, Borovskii and ~archukova'~ have stated that the determination of the solubility limits of phase diagrams using high-resolution micro-analyzer measurements at the interfaces of multiphase couples is not an accurate technique because of deviations from the equilibrium compositions at a moving interface; diffusion couples may be used to map out the phase boundaries in the equilibrium diagram, but the final determination of the solubility iimits should be made with equilibrated samples. The purpose of this work was to investigate the conditions prevailing at an interface in a multiphase diffusion couple and to compare the interface compositions with those associated with true thermodynamic equilibrium between the two phases. Microanalyzer techniques were used to measure interface compositions in two-phase Cu-A1 diffusion couples annealed at 80@, 905", and 1000°C for various times.

Jan 1, 1969

By C. S. Roberts

The creep mechanism and kinetics of fine-grained magnesium have been studied over the temperature range 200&apos; to 600°F. As a result of a photographic study of microstructural changes, transient and steady-state creep components have been correlated with slip, subgrain formation, and cyclic deformation at the grain boundaries. THE approach of this research has been the blend of a quantitative study of the creep strain of polycrystalline magnesium as a function of time, stress, and temperature with direct microstructural observations of the operative deformation processes. The validity of the conclusions is dependent on the condition that the microstructural changes seen on the polished surface qualitatively represent those occurring in the bulk of the metal. The work was intended as much to lay a background to a study of highly creep-resistant magnesium alloys as to provide a description of the behavior of the base metal itself. The spectroscopic analysis of the electrolytic magnesium used in this study is as follows: Al, 0.009 pct; Ca, <0.01; Cu, 0.0011; Fe, 0.021; Mn, 0.012; Ni, 0.0004; Pb, 0.0012; Si, <0.001; Sn, <0.001; and Zn, <0.01. The impurity level is approximately that of commercial magnesium alloys. The original ingot was melted under Dow type 310 flux and cast as a 3 in. diam billet. It was extruded into 1 in. flat stock under the conditions: billet preheat 800°F (1 hr), container and die temperature 800°F, speed 3 ft per min, and area reduction ratio 45:1. The extrusion process was chosen in preference to rolling and recrystallization because it allowed easier grain size control from specimen to specimen. The grains of the extruded metal were fairly equi-axial and uniform in the size range of 4 to 6 thousandths of an inch. The preferred orientation of basal planes about the transverse direction was determined by an X-ray diffraction surface reflection method. A beam of filtered copper radiation was directed at an angle of 17" to both the transverse direction and the surface yet perpendicular to the extrusion axis. Analysis of the (002) diffraction arcs in the resulting photographic patterns gave an approximate intensity distribution along the great circle which extends through the center of the basal plane pole figure and to the extrusion axis poles. Successive layers of metal were removed by macro-etching between exposures. The extruded texture is relatively sharp, but the most significant point is the position of the maximum basal plane pole density and its variation with depth below the surface. Fig. 1 shows that this maximum is rotated 15" from the normal at the surface toward the extrusion direction. Such an inclination has been reported for extruded 1 pct Mn and 8 pct A1-0.5 pct Zn alloys.&apos; The inclination decreases until the maximum splits at about 0.025 in. depth into two elements of equal and opposite rotations from the ideal. The double texture persists to as great a depth as was experimentally convenient to examine. It probably continues to the very center of the extrusion. There is no great change in the sharpness of the individual elements of the texture with depth. A plate of metal about 0.015 in. thick at the surface of the extruded stock was produced by etching. A transmission diffraction pattern was made for the purpose of determining any preferred orientation of a direction in the basal planes. Relatively uniform {loo) and {101) rings were produced. There is little tendency for parallelism of a given direction in the plane with the projection of the extrusion axis on it. The creep specimens were machined from 6¼ in. lengths of the extruded stock. Creep was measured on the reduced section, ½x1/8X2¼ in. long. This section was electropolished on one side for the studies of microstructural changes during creep. An orthophosphoric acid-ethyl alcohol electrolyte was used under the conditions recommended by Jacquet.² Hand polishing was used for previous mechanical preparation. Electropolishing was continued until all mechanical twins had been removed. The electro-polished surface was protected from oxidation during creep testing by a thin layer of silicone oil. All micrographs were taken at room temperature on conventional metallographic equipment and after removal of the oil film. The creep tests were performed with machines which have been described in detail by Moore and McDonald." Five testing temperatures, 200°, 300°, 400°, 500°, and 600° ±3°F were used. Difference in temperature between the two ends of the specimen reduced section was 2°F or less. The testing was done at constant load. Strain readings were taken as frequently as necessary to develop usable creep curves. Tensile Creep vs Time, Stress, and Temperature A definition of terms is necessary. Whenever successive sections of a creep strain-time curve show decreasing, constant, and increasing slope with time they will be termed primary, secondary, and tertiary

Jan 1, 1954

By Benny Langston, Frank M. Stephens

Although little information is available concerning small-scale equipment for dust collection in laboratories, it is possible to adapt standard equipment for laboratory use. Dust from laboratory processes may be collected by cyclone separators, filters, electrostatic separators, scrubbers, and settling chambers. IN recent years much attention has been given to recovery, treatment, and disposal of dusts discharged into the atmosphere from operations of industry. considerable data has been accumulated on both operation and design of dust-collector equipment for commercial installations. On the other hand, there is almost no published data on design and construction of small-scale equipment to handle dust problems that arise in the ore-dressing laboratory. Dust-collection equipment such as multiclones, single-cyclones, scrubbers, chemical and mechanical filters, settling chambers, and electrostatic separators has proved its efficiency for collecting dust in industry. In the laboratory, however, the engineer is faced with the problem of collecting small quantities of dust, inexpensively, without diverting the major effort from the metallurgical problem to the problem of collecting dust produced by the process. For most applications standard dust-collection equipment is too large for use in the laboratory; however, for control of dust in large working areas it is often possible to use a standard dust collector, such as an air filter, with branch ducts to eliminate a health hazard. For example, the well-furnished sample-preparation room containing small jaw crushers, rolls, and pulverizers, in addition to the riffles and screens necessary for preparation of samples, presents a perennial source of dust. The authors&apos; experience has shown that a combination system consisting of overhead branch ducts to the individual equipment and floor ducts with grills, where applicable, connected to a central dust collector effectively removes dust generated in preparation of samples. Fig. 1 is a sketch of a downdraft dust-collector for table installation. Similar systems can be built with floor grids. For portable equipment such as laboratory vibrating screens this type of installation with a steel grill to support the heavy load is reasonably efficient. Overhead branch ducts to individual crushing and grinding equipment, although efficient, must be carefully controlled by dampers to prevent excess loss or a change in the composition of the sample. Change in sample composition can result from excess velocity, which causes selective removal of constituents of low specific gravity. Fig. 2&apos; shows the theoretical effect of terminal velocity on spherical particles of different specific gravities in air and water under action of gravity. Fig. 3 shows the effect of air velocity on composition of CaCO, coal mixtures. Although the entrainment of dust particles in a moving air stream is the basic mechanism by which all dust-collection equipment functions, usually intake velocity of the dust-collection system must be controlled to prevent loss of part of the sample. As an example of what may happen when excess velocities are used, a mixture of 50 pct coal and 50 pct limestone was crushed to —10 mesh and fed to a pulverizer equipped with an overhead dust-collection system. Fig. 4 shows the overhead dust-collection equipment used in this test. The pulverizer was set to give a product 95 pct —100 mesh in two stages. Velocity of air passing over the lip of the pulverizer was measured with an anemometer. After grinding, the finished product was analyzed to show the amount of calcium carbonate present. Fig. 3 shows graphically the increase in calcium carbonate as velocity through the dust-collection duct was increased. These data show that at a velocity of 1 ft per sec little if any of the coal was entrained by the overhead draft. At the maximum velocity, about 6.5 ft per sec, approximately 7 pct more coal was entrained than calcium carbonate. From an operating standpoint, this problem can be remedied easily by dampering the line to control velocity. The lowest velocity commensurate with satisfactory dust control should be used to prevent excess loss and, in some cases, selective dust loss. Collection of Dust in Laboratory Processes As in industry, the engineer desires to collect efficiently the dust produced by processes being investigated on a laboratory scale. However, in the collection of laboratory dusts he is faced with two additional problems: 1—The volumes of gas and the quantity of dust that must be recovered are small when compared with the capacity of standard dust-collector equipment, which must be scaled down in design except for collection of dust from large pilot-plant operations. 2—In addition, because of the variety of problems studied in the process laboratory, the engineer cannot design today a dust collector that will meet the conditions imposed by the processes of tomorrow. Sometimes, therefore, he must compromise collection efficiency to minimize the cost of fabrication and the amount of time diverted from the metallurgical to the dust-control problem. For collection of dust from laboratory processes a cyclone separator, filters, electrostatic separators, scrubbers, and settling chambers can usually be adapted for small-scale operations. The following

Jan 1, 1955

By Erhard M. Winkler

The rapid decay and disfiguring of stone monuments in urban and desert rural areas has challenged conservators to protect stone surfaces from premature decay. They attempt to halt the natural process of stone decay and possibly to restore the original strength lost mostly by chemical weathering and the loss of binding cement. Ageneral solution is not possible because the physical and chemical characteristics must be considered for different stone types. The failures of stone preservation and restoration are greater in number than the cures. The need for repair of stone decay goes back to evidence of Roman replacement of decaying stone. The presence of excess water in buildings has long been recognized. Moisture tends to enter masonry from air in humid climates, a most important but often underrated factor (Fig. 1) suggesting that sealing should be the answer. Undesirable staining and efflorescence result in accelerated scaling. Today, the great variety of chemicals available to the modem conservator for sealing. consolidating, or hardening stone fall into two very different categories: surface sealers and penetrating stone consolidants, or a combination of both. SEALERS Sealers develop a tight, impervious skin which prevents access of moisture. Surface sealing has saved monuments from decay by eliminating the access of atmospheric humidity. Pressure tends to develop behind the stone surface by moisture escape. Efflorescence, crystal growth action, and freezing can cause considerable spalling (Anderegg, 1949). Flaking results when moisture is trapped behind the sealed surface. Yellowing and blotchiness are also frequently observed. The following sealants are in common use today: linseed oil, paraffin, silicone, urethane, acrylate, and animal blood on stone and adobe. Extensive cracking and yellowing has resulted soon after application. In the past many such treatments have created more problems than cures: 1. Linseed oil and paraffin have been in use for centuries. Embrittlement and yellowing occur rapidly because these are readily attacked by solar ultraviolet radiation. 2. Animal blood as paint has temporarily waterproofed adobe mud and stone masonry. The origin of blood paint has a religious background rooted in the Phoenician and Hebrew cultures. Instant water soluble dried blood can substitute for fresh blood. Winkler (1956) described the history and technique of the use of blood. 3. Silicones have proven very effective and are long lasting. In contrast, acrylates, urethane, and styrene are generally rapidly attacked by UV radiation (Clark et al., 1975). Sealing of Different Rock Types Granitic rocks have a natural porosity traced to 4.5% contraction of quartz, during cooling of the parent magma, compared with only 2% contraction of all other minerals; protection against the hygric forces may require waterproofing of granite in some in- stances. The Egyptian granite obelisk in London is an example. Soon after its relocation from Egypt to London, Cleopatra&apos;s Needle was treated, in 1879, with a mixture of Damar resin and wax dissolved in clear petroleum spirit; surface scaling became evident after half a year of exposure to the humid London atmosphere. The treatment of the ancient granite monument from Egypt has denied access of high relative humidity (RH) in London to the trapped salts inherited from the Egyptian desert and has protected the monument from decay (Burgess and Schaffer, 1952). The sister obelisk set up in Central Park, New York City, has fared less favorably because similar treatment was done too late, only after the salts hydrated and hundreds of kilograms of scalings disfigured the obelisk surface (Winkler, 1980). Surface coating of other common stones may be needed. Crystalline marble absorbs moisture from high RH atmospheres: dilation may ensue when curtain panels bow as the moisture starts to expand during daily heating-cooling cycles. A good sealer may prevent the moisture influx provided that no moisture can enter from the inside of the building. Limestones, dolomites and all carbonate rocks are subject to dissolution attack by rainwater, especially in areas where acid rain prevails (Fig. 2). The interaction of sulfates in the atmosphere with the stone can be halted by waterproofing to avoid the formation of soft and more soluble gypsum. The stone surface attack can be diminished if nearly insoluble Ca-sulfite crusts can form, instead of Ca-sulfate. Replacement of fluorite or barium compounds at the stone surface acts as a hardener, rather than a sealant. Sandstones have generally high porosity and rapid water travel can occur along unexpected routes and from any direction. Any surface sealing may do more damage by scaling and bursting than if the stone is left without treatment. Sealing of sandstones is therefore not advised at any time. Testing the efficiency of sealants: Several authors discuss waterproofing materials, silicones, urethanes, acrylates and stearates, as to their water absorption, spreading rates of water on the treated surface, water vapor transmission, resistance to efflorescence, and general appearance (Clark et al., 1975). De Castro (1983) measured the angle of contact of a microdrop (0.004 cm3) on a stone surface as characteristic of the wettability. Laboratory tests and limited field performance are described by Heiman (1981). The crest of a Gothic sandstone arch, which was sealed with silicone,

Jan 1, 1994

By R. T. Hukki

John F. Myers (Consulting Engineer, Greenwich, Corm.)—Since the art of comminution has lain practically dormant for many years, it is very interesting that R. T. Hukki approaches the subject with a new concept. One is reminded of the research carried on by A. W. Fahrenwald of Moscow, Idaho, a few years ago. Fahrenwald mounted a steel bowl on a vertical shaft. The balls and ore placed in the bowl were rotated at fast speeds, thus simulating the supercritical speeds used by Hukki. The rolling action of the balls against the smooth shell liner has pretty much the same effect. The action is horizontal in one case and vertical in the other. Both researchers report good grinding activity. It is also constructive that such able investigators give to the students of comminution their interpretation of their laboratory results in terms of large-scale operation. History shows that it takes a lot of time for such radically new ideas to be absorbed by the industry. Typical of this is the present-day activity of cyclone classification in primary grinding circuits. The idea of cyclone classification has been kicking around for 30 or 40 years. Certainly we all suspect that the ponderous grinding mills of today, and their accessory apparatus, large buildings, etc., will ultimately give way to small fast units, just as this has occurred in other industries over the past 50 years. At the moment there is no evidence that ball and liner wear is prohibitively high. In fact, at the time Fahrenwald was demonstrating his high-speed horizontal machine at the meeting of the American Mining Congress, several years ago, he assured this writer that the balls retained their shape much longer than they do in conventional tumbling mills. Rods and balls that slide (as some operators in uranium plants are experiencing) get flat. Apparently the balls have a rolling action. Mr. Hukki&apos;s references to the processing capacity of the Tennessee Copper Co. mills is adequate. Those studying this subject will be greatly interested in the paper presented by Richard Smith of the Cleveland-Cliffs Iron Co. at the annual meeting of the Canadian Institute of Mining and Metallurgy in Vancouver April 24, 1958. This paper will be published during the latter part of 1958 in the Canadian Institute of Mining and Metallurgy Bulletin. Hukki&apos;s pioneering spirit is to be commended. R. T. Hukki (author&apos;s reply)—It has been heartening to read the objective discussion by J. F. Myers. The sincerity of his opinions is further strengthened by the fact that the article he has discussed contradicts in a major way the parallel achievements of his life work. Myers is right in his opinion that in general it takes a long time before new ideas are accepted by the industry. On the other hand, revolutions usually take place at supercritical speeds. There are many indications at present that both the unit operation of grinding and the related subject of size control are now just about ripe for a revolution. In grinding, brute force must ultimately give way to science. Rapid progress can be anticipated in the following fields: 1) Autogenous fine grinding at supercritical speeds will be the first advance and the one that will gain recognition most easily on industrial scale. At this moment, little Finland appears to be leading the world. Crocker recently made a statement that in nine cases out of ten, your own ore can be used as grinding medium more effectively and far more economically than steel balls. This is true. The present author would like to introduce a supplementary idea: In eight cases out of the nine cited above, it can be done at the highest overall efficiency in the supercritical speed range. Fine grinding must be based on attrition, not impact. The path of attrition may be vertical, horizontal, even inclined. 2) In coarse grinding, the conventional use of rods is sound practice. However, even the rods can be replaced by autogenous chunks large enough to offer the same impact momentum as the rods. To obtain the momentum, the chunks must be provided with a free fall through a sufficient height in horizontal mills operated at supercritical speeds. Coarse grinding must be based on impact. Detailed analysis of the subject may be found in a paper entitled "All-autogenous Grinding at Supercritical Speeds" in Mine and Quarry Engineering, July 1958. 3) All conventional methods of classification, including wet and dry cyclones, are inefficient in sharpness of separation. Continuous return of huge tonnages of finished material to the grinding unit with the circulating load is senseless practice. In the near future the present methods will be either replaced or supplemented by precision sizing. These three fields are also the ones to which J. F. Myers has so admirably contributed in the past. Fine Grinding at Supercritical Speeds. By R. T. Hukki (Mining EnGineERInG, May 1958). Eq. 9, page 588, should be as follows: T , c, (a — 6&apos;) n D Ltph On page 584 of the article the captions for Figs. 4 and 5 have been placed under the wrong illustrations and should be interchanged.

Jan 1, 1959

By E. Klar, A. B. Michael

Differences in the electrical conductivities of copper powder sintered under reducing, selectively oxidizing, and neutral atmospheres are related to impurities in solution or as precipitated oxides. The precipitation of impurities as oxides during sintering in nitrogen is proposed for maximizing the conductivity of sintered copper. Conductivity equations for two-phase systems are summarized. Selected equations are applied to porous sintered copper and composite structures. A recent review of the influence of impurities on the electrical conductivity of copper by Gregory et al.1 emphasized that an impurity in solid solution has a much more pronounced effect on reducing the electrical conductivity than when present partly or wholly as a second phase. When impurities in solid solution can be precipitated as oxides, the copper is purified with respect to these elements and the conductivity is increased. Cast and wrought copper, therefore, frequently contain an intentional residual oxygen concentration to oxidize and precipitate impurities less noble than copper. Copper powders generally also contain impurities which can contribute to a reduction in the electrical conductivity of the sintered material. The most deleterious impurities commonly found in commercial copper powders which markedly decrease the electrical conductivity when in solid solution but which can be precipitated as oxides include iron, tin, antimony, arsenic, cobalt, and nickel. In addition to impurities, porosity in sintered copper also contributes to a reduction in the electrical conductivity. The work reported herein discusses the influence of impurities, sintering atmospheres, and porosity on the electrical conductivity of sintered copper. These are important factors for controlling the electrical conductivity of materials such as sintered copper electrical contacts. Several publications on the electrical conductivity of sintered copper and composite materials with random pores or obstacles have incorrectly considered the conductivity to be proportional to the volume fraction of conducting material. However, analyses and equations have been proposed, the earliest of which is perhaps Lord Raleigh&apos;s of 1892,&apos; which more accurately describe the conductivity of two-phase systems. These equations which in many cases allow one to closely estimate the electrical conductivity of porous sintered materials and two-phase composites will be reviewed and related to the measured electrical conductivity of sintered copper, copper-graphite. and Ag-W composites. INFLUENCE OF IMPURITIES AND OXIDIZING, REDUCING, AND NEUTRAL ATMOSPHERES Zone-Melted and Leveled Copper. The electrical conductivities of an electrolytic copper and a lower-purity compacting grade of copper powder after consecutive treatments in reducing or selectively oxidizing atmospheres are compared in Fig. 1. The powder and drillings from the electrolytic copper ingot were melted and solidified in a graphite crucible under hydrogen. The vertical zone leveling technique of multiple passes in opposite directions was used to obtain a uniform distribution of impurities. The electrical conductivity was measured on machined specimens 3 by 0.10 in. diam of the zone-leveled material and after the same specimens were consecutively treated as follows: 1) heating in hydrogen at 1600°F for 3 hr; 2) heating in air in a sealed tube so that 0.04 pct 0 was introduced; heating was started at 1000°F for 1 hr and continued at 1800°F for 8 hr; the stepwise diffusion treatment was used to avoid loss of copper due to evaporation of copper oxide at higher temperatures; and 3) final heating in hydrogen at 1600°F for 3 hr. A L&N Kelvin Bridge, type 4306, and a test fixture with knife edges 2 in. apart were used to measure the room-temperature conductivity with an estimated accuracy of ±0.5 pct. In all cases the conductivity of the electrolytic copper was approximately 100 pct of IACS. The conductivity of the impure copper was 80 pct of IACS both in the cast state and after heating in hydrogen. However, after the heating in air, the conductivity of the impure material was 100 pct IACS. After again heating in hydrogen, the electrical conductivity decreased to 60 pct of IACS. This decrease is attributed both to the dissolution of impurities and observed intergranular cracks due to the phenomenon of hydrogen embrittle-ment in copper. These data show that the electrical conductivity of commercial copper as represented by a compacting type powder can be increased significantly by the precipitation of impurities as oxides through heat treatment in an oxidizing atmosphere. Sintered Copper Powder. A commercial compacting type of copper powder pressed to various densities was sintered either in nitrogen, dissociated ammonia, or dissociated ammonia followed by a selectively oxidizing atmosphere of air in sealed Vycor tubes so that 0.06 pct O was added to the material. The electrical conductivity was measured perpendicular to the pressing direction on as-sintered specimens of approximately 3 by 0.25 by 0.15 in. The fully dense material was obtained by zone melting and leveling

Jan 1, 1969

By E. W. Felegy

THE need for improved mine communication to increase efficiency and to insure greater safety has long been recognized. General and unrestricted communication between all points underground, and between the surface and all points underground, is probably the least advanced phase of the mining industry. An ideal system of mine communication must require no fixed wire installations. The equipment must be small, lightweight, and readily portable, and the power requirements low. A system that can be used not only under normal circumstances but also in an emergency, when the continuity of wires, tracks, and pipelines may be disrupted, must function independently of any aid furnished by standard installations. Radio communication offers possibilities of meeting all the requirements necessary for an ideal communication system in underground mines. Transmission of signals must be achieved through one or both of two mediums, through the air in mine openings or through the strata. The results or lack of results obtained by early investigators showed conclusively that radio communication by space transmission cannot be accomplished effectively beyond line-of-sight distances in underground passageways. A radio system underground therefore must depend solely upon transmission through soil and strata. The application of radio to underground mine communication was investigated by many individuals and agencies at different times in the last several decades, but little success was achieved before World war 11.2-0, The results of experiments during the war, and further knowledge gained in experiments with vastly improved communication methods and equipment after the war provided the background for additional research in radio communication in underground mines. During 1950 to 1.952 the University of Minnesota sponsored an investigation to determine the possibility of developing: a system of radio communication universally applicable in underground metal mines in the Lake Superior district. The possibility of using radio equipment to determine the imminence of rock bursts in deep copper mines in that district also was investigated. The investigation supplemented previous and concurrent emergency mine communication studies of the U. S. Bureau of Mines. Testing equipment and laboratory facilities maintained by the Bureau of Mines at Duluth, Minnesota, were used in the research program, which was conducted as a mining engineering graduate research problem by the present writer under the direction of T. L. Joseph and E. P. Pfleider. The radio units used in the research program were designed and built to specification solely to conduct tests of radio communication in mines. Two identical units were used in all tests. Each unit contained a transmitter section, a receiver section, and a power-supply section, mounted on a single chassis. The entire unit was enclosed in a single 10x12x18-in. metal case provided with a leather-strap handle for carrying purposes. The front of the case was hinged to open upward and provide easy access to the single control panel on which all controls were mounted. Storage batteries supplied the operating power for all tests. Standard 6-v automobile batteries were utilized to provide adequate capacity to conduct tests for a full day without exhausting the battery. A frequency range from 30 to 200 kc was covered in eight pre-fixed steps on each unit. The carrier frequencies were crystal-controlled and amplitude-modulated. The receiver employed an essentially standard superheterodyne circuit and was sufficiently sensitive to detect signal strengths of 5 micro v. A heterodyne circuit was employed in the transmitter to obtain the low-carrier frequencies used in the units. Power output of the transmitter, usually less than 2 w, rarely exceeded 3 w in any test. Tests were conducted in mines on the Vermillion iron range in Minnesota, the Gogebic iron range in Wisconsin, the Menominee and Marquette iron ranges in Michigan, and a copper mine in the upper Michigan peninsula. All tests were conducted when the mines were operating normally, and usual mining, maintenance, and transportation activities were in progress, so that any interference caused by normal production activities could be evaluated during the tests. Tests were made between different points underground in each mine, and between underground and surface points at some mines. Test readings obtained at any one mine were calibrated in the laboratory before another series of tests were begun at the next mine. The transmitter and receiver were separated by one or more levels in each test, and generally there was no other means of communication between test points. Two 100-ft lengths of rubber-covered wire were used for antenna wires on each unit in both transmission and reception. The ends of the wires were connected to ground points in one of several methods, depending upon physical conditions at each test site. The wires were clipped to metal rods about 200 ft apart in the back, side, or bottom of the mine opening where the character of the rock permitted driving rods. Both wires were clipped to points about

Jan 1, 1954

By H. E. Mauck

The many factors that control bumping must be carefully studied for each coal seam where bumps occur, and specifications known to exclude bumping should be incorporated in the mining plans. This calls for complete knowledge of the seam&apos;s characteristics and its adjacent strata, and in many instances these characteristics are not revealed until the seam is actually mined. Pressure and shock bumps, the two general types, occur jointly and separately. In this discussion no differentiation will be made. Whether pressure or shock, they are treated as bumps, and both must be eliminated. Bumps in mines have occurred in several places throughout the coal fields of the world. A study of many of these occurrences indicates that geologic characteristics, development planning, and mining procedure have contributed. But more specifically, there are conditions usually associated with bumps: thickness of cover, strong strata directly on or above the seam, a tough floor or bottom not subject to heaving, mountainous terrain, stressed and steeply pitching beds, and the proximity of faults and other geologic structures. Mine planning should incorporate these known factors (not necessarily in order of importance): 1) Main panel entries should be limited to those absolutely necessary to ventilate and serve the mine. This reduces the span over which stresses may be set up that will later throw excessive pressures on barrier and chain pillars when they are being removed. 2) Barrier pillars should be as wide as practicable so that they will be strong enough to carry the loads thrown on them when final mining is being carried out. 3) Pillars should never be fully recovered on both sides of a main entry development if the barrier and chain pillars are to be removed later. The excessive pressures placed on the main chain and pillar barriers by arching of the gob areas can result in bumping when these barriers are being removed. 4) Full seam extraction is better accomplished by driving to the mine boundary and then retreat-drawing all pillars. If there are natural boundaries in the mine—such as faults, want areas, and valleys —retreat should be started there. 5) Pillars should be uniform in size and shape. The entire development of the mine should call for uniform blocks with entries driven parallel and perpendicular. Only angle break-throughs should be driven when necessary for haulage, etc. 6) For better distribution of rock stresses and reduction of carrying loads per unit area, both chain and barrier pillars should be developed with the maximum dimensions. 7) Pillars should be open-ended when recovered. If they are oblong, the short side should be mined first. Both sides of a block should not be mined simultaneously, but under no circumstance should the lifts be cut together. 8) Pillar sprags should not be left in mining. If they are not recoverable, they should be rendered incapable of carrying loads. 9) Pillar lines should be as short as practicable. (Three or four blocks are adequate). Experience has shown that rooms should be driven up and retreated immediately. The longer a room stands, the more unfavorable the mining conditions. This contributes to bumping. 10) Pillars should not be split in abutment zones (high stress areas lying close to mined out areas) and if slabbing is necessary, it should be open-ended. 11) Pillars should be recovered in a straight line. Irregular pillar lines will allow excessive pressures thrown on the jutting points. Experience has shown that the lead end of the pillar line can be slightly in advance. 12) Pillar lines should be extracted as rapidly as possible. This appears to lessen pressures on the line and render abutment zones less hazardous. 13) Extraction planning should call for large, continuous robbed out areas. Robbing out an area too narrow to get a major fall of the strata above the seam tends to throw excessive pressures on a pillar line. 14) Timbering in pillar areas should be adequate but not excessive. Too heavy timbering or cribbing is likely to retard roof falls and throw excessive weight on the pillar line. 15) Experience has shown that when pillar lines have retreated 800 to 1000 ft from the solid, bumps can occur. Because this distance may vary in different seams, impact stresses should be studied for each individual condition. In any event, extra precautions should be taken against bumps in this area. This list of controlling factors may or may not be complete. It probably is not, but it covers most of the problem&apos;s significant aspects. The question is whether or not bumping can be eliminated. The answer is that bumping can be minimized and possibly eliminated if these and other established factors are thoughtfully considered and incorporated in the mining and extraction plans. If a mine has already been developed or the pattern set so that little change can be made, then it will be necessary to adjust to the most nearly practicable system that can incorporate the known factors.

Jan 1, 1959

By W. M. Robertson

The formation of grain boundary grooves on surfaces of poly crystalline samples of cobalt and iron immersed in liquid lead has been studied. The grooves form by volume diffusion of the solutes cobalt and iron in the liquid. The diffusion coefficients of the solutes in liquid lead are derived from the measured rate of grooving. The diffusion coefficients are described by the relation D = Do exp (-Q/RT), with, for cobalt, Do = 4.6 x 10-4 sq cm per sec and Q = 5300 ± 800 cal per mole, and for iron, DO = 4.9 x 10-3 sq cm per sec and Q = 10,500 ± 1500 cal per mole. LIQUID metal-solid metal interactions occur at solid-liquid interfaces. Interfacial energy provides a driving force to change the morphology of the interface. Mullins1,2 has derived expressions for the kinetics of interface morphology changes driven by capillarity. These expressions can be applied to an isothermal system of a solid in equilibrium with a liquid saturated with the solid. Surface profile changes can occur by volume diffusion of the solute in the liquid, by volume self-diffusion in the solid, and by interfacial diffusion at the liquid-solid interface. A groove will form at the intersection of a grain boundary with a solid-liquid interface, reducing the total interfacial free energy of the system. The solid-liquid interfacial energy ? must be greater than half the grain boundary energy of the solid ?6 for Mullins&apos; calculations to apply. If ? is less than ?b/2, then the liquid penetrates the boundaries, separating the grains rather than forming grooves. Boundary penetration did not occur in the work described here. where CO is the equilibrium volume concentration of the solid in the liquid, Dv the volume diffusion coefficient of the solid in the liquid, ? the interfacial free energy of the solid-liquid interface, O the atomic volume of the solid crystal, k Boltzmann&apos;s constant and T the absolute temperature. Eqs. [1] and [2 ] also apply to grooving by volume self-diffusion in the solid,1 with CoODv = D Self, where DSelf is the volume self-diffusion coefficient of the solid. For a grooving mechanism of interfacial diffusion at the solid-liquid interface, the groove width is given by2 where CS is the interfacial concentration of the diffusing species, and DS is the interfacial diffusion coefficient. Eqs. [1] and [3] can be used to determine the mechanism of groove growth. A t1/3 dependence of the growth indicates volume diffusion and t1/4 indicates interfacial diffusion. In some cases, volume diffusion and interfacial diffusion both can contribute substantially to the grooving process, causing the time dependence to be intermediate between t 1/3 and t1/4.3 For these cases, the relative contributions of the two processes can be separated.4 However, in many cases, one process will be dominant, and the data can be analyzed on the basis of Eq. [1] or Eq. [3] alone. The time dependences for volume diffusion in a liquid and volume self-diffusion in a solid are the same. However, the self-diffusion contribution of the solid is usually negligible compared to volume diffusion in the liquid. After the grooving mechanism has been determined, Eq. [1] or Eq. [3 ] yields the kinetic parameter A or B. The kinetic parameter can be used to calculate values for the unknown quantities in the product CD?. Usually C is known or can be estimated. If ? is known, then D can be calculated. In a measurement of grain boundary grooving of copper in liquid lead,&apos; the time dependence indicated volume diffusion in the liquid. The quantities Co, Dv, and ? were obtained from the literature, giving excellent agreement between the observed values of A and the values calculated from Eq. [2 ].5 In a study of the grooving of several refractory metals in liquid tin and liquid silver, A1len6 educed that grooves formed by volume diffusion in the liquid. In a study of nickel in a nickel sulfide melt, Steidel, Li, and spencer7 found volume diffusion grooving kinetics. Both Dv and ? were unknown, so they could not obtain either one separately, though they did obtain a reasonable value for the temperature dependence of the product Dv ?. Several methods have been used to obtain surface profiles. It can be done by sectioning through the interface7 or by chemically removing the liquid from the solid surface after solidification of the liquid.6 However, if the liquid dewets the solid on removing the solid from the melt, then the interface can be observed directly. This method was used previously&apos; and was utilized also in the present study. EXPERIMENTAL PROCEDURE Lead of 99.999 pct purity was obtained from American Smelting and Refining Co. Cobalt sheet was obtained from Sherritt-Gordon Mines, Ltd., with a nominal purity of 99.9 pct, the principal impurities being nickel, iron, copper, carbon, and sulfur. The sheet was

Jan 1, 1969

By D. J. Carney, R. L. Stephenson

A sampling technique has been developed for procuring a sample of sinter representative of the entire depth of the sintering bed. The sampling method involves the use of an open-bottom metal basket that rides on the grate of the sintering machine and when removed contains a sample of the sintered product. Additional data have been obtained to indicate that the tumbler test is a suitable means of measuring sinter strength. IN the last few years additional sintering facilities have been installed in both the Pittsburgh and the Chicago district of the United States Steel Co. Since the construction of these sintering plants made possible the use of higher percentages of flue-dust sinter in our blast-furnace burdens, it became important to study means of controlling the quality of sinter to obtain optimum results in the blast furnace. For controlling an operating process, it is necessary first to establish standards by which the quality of the product can be judged. For sinter, it appeared that an important property was its strength or its resistance to degradation during transportation and charging into the furnace. Consequently work was undertaken to establish a standard for sinter strength that could be used both for controlling sintering-plant operations and for correlating sinter quality with blast-furnace performance. The first problem in setting up a standard was that of procuring a sample that would be representative of the sinter made under any particular set of conditions at the sintering plant. Since the United States Steel Co. sintering plants discharge the finished sinter either into a large pit or onto a rotary cooler, the sinter becomes inseparably mixed with material sintered 2 hr before or 2 hr afterwards. For this reason the exact identity of the sinter is lost. A sample selected as the cooler is discharged, or as the sinter is removed from the pit, cannot be said to be truly representative of the sinter made at any specific time. Sampling The first attempt to procure a sample that would be representative of a specific sinter mix and of specific operating conditions was made by stopping the Dwight Lloyd sintering machine and removing an entire pallet full of sinter. This method, however, proved very difficult to perform and interfered considerably with the operation of the plant. To overcome this difficulty, a sampling method was devised by technologists at South Works enabling them to secure, without interrupting the sintering operation, a sample of about 1 cu ft of sinter, representative of sinter for the full depth of the sintering bed. The South Works method involves the use of a steel-frame-work basket. A typical basket is shown in Fig. 1. The basket has been used both with and without crossbars along the bottom. As long as the crossbars are in the same direction as the grate bars on the sintering machine they do not interfere with the sintering process. The basket is set on an empty grate of the Dwight Lloyd sintering machine before it passes under the swinging feed spout, see Fig. 2. When the basket is removed after it has travelled the length of the sintering machine, it contains the sample. Just before the basket is removed, the sinter is scored and chipped to facilitate removal of the sample from the sinter bed. A view of the basket after its removal is shown in Fig. 3. Although the sampling method was originally designed for use on a Dwight Lloyd sintering machine, it can also be used on the Greenawalt type of machine. When used on the Greenawalt-type machine, the basket is placed on the sintering grate before the charging car passes over it, and finally it is removed just before the pan is dumped. Testing After a method of obtaining a representative sample of sinter had been developed, the next step was to select a method of measuring its strength. The irregular shape and size of the sinter pieces precluded the use of a simple compression test for determining strength; consequently, the shatter test and tumbler test were investigated. To perform the shatter test, a sample of sinter, approximately 5 lb, is dropped from a hinged-bottom box at a height of 3 ft onto a steel plate. The broken sinter is sieve-analyzed after a specified number of drops. The tumbler test is performed with the use of a standard ASTM coke-tumbling drum. The drum is 3 ft in diam and is equipped with two lifter bars diametrically opposite one another on the inner periphery of the drum. The drum is rotated at a speed of 24 rpm for 200 revolutions, and after tumbling the sample is sieve-analyzed. To express as single numbers the results of sieve analyses after shattering or tumbling, the method suggested by R. E. Powers1 was employed. This method involved plotting the size of the sieve openings on a logarithmic scale and the cumulative per cent larger than each sieve on a probability scale as described by J. B. Austin.&apos; By interpolating from the plotted data, which in most cases approximated

Jan 1, 1954

By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960&apos;s. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60&apos;s. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981

By J. P. Hammond, C. J. McHargue

The wire textures for cold rolled and recrystallized iodide titanium and the sheet textures for this material produced by cold and hot rolling, and recrystallization at a series of temperatures were determined. &apos;The effect of the a + ß transformation on the sheet texture was noted. UNTIL recently it was believed that all hexagonal close-packed metals deformed by slip on the basal plane, (0001), and that rolling should tend to rotate this slip plane into the plane of the rolled sheet. The pole figures of cold rolled magnesium&apos; are satisfactorily explained on this basis. There is a tendency for the <1120> directions to align parallel to the rolling direction, and the principal scatter is in the rolling direction. Zinc% as a rolling texture in which the hexagonal axis is inclined 20" to 25" toward the rolling direction. Twinning is believed to account for the moving of the basal plane away from parallelism with the rolling plane. The texture of beryllium3 places the basal plane parallel to the rolling plane with the [1010] direction parallel to the rolling direction, and the scatter from this orientation is primarily in the transverse direction. Cold rolled textures reported for zirconium&apos; and titanium5 how the [1010] directions to lie parallel to the rolling direction and the (0001) plane tilted by approximately 25" to 30" to the rolling plane in the transverse direction. Rosi has recently reported that the mechanisms for deformation in titanium are distinctly different from those commonly reported for hexagonal close-packed metals. The principal slip plane is the prismatic plane, {1010), with some slip also occurring on the pyramidal planes, (1011). However, there is no evidence for basal slip. The slip direction is reported to be the close-packed digonal axis, [1120]. In addition to the twin plane commonly reported for metals of this class, {1012), Rosi found the twin planes (1122) and {1121), with the dominant twin plane being (1121). Information regarding the recrystallization and hot rolling textures of hexagonal close-packed metals is limited. Barrett and Smigelskas report that rolling beryllium at temperatures up to 800°C and recrystallization at 700°C produce textures not differing from the cold rolled sheet texture.3 McGeary and Lustman find that hot rolling at 850°C produces the same basic texture in zirconium as rolling at room temperature.&apos; These investigators also report that the texture for sheet zirconium recrystallized at 650 °C differs from the cold rolled orientation inasmuch as the [1120] direction, instead of the [1010] direction, is parallel to the rolling direction. In the case of titanium, it is not possible to deduce which direction is preferred in the recrystallized state from the pole figures presented by Clark." The purpose of this paper is to report an extensive investigation of the preferred orientations in iodide titanium. Since the deformation mechanisms for titanium are different from those commonly given for hexagonal close-packed metals, it is not surprising to find distinct differences between the textures of titanium and other metals of this class. Materials and Methods This investigation was carried out on iodide titanium obtained from the New Jersey Zinc Co. with an analysis as follows: N2, 0.002 pct; Mn, 0.004; Fe, 0.0065; A1, 0.0065; Pb, 0.0025; Cu, 0.01; Sn, 0.002; and Ti, remainder. The crystallities of titanium were broken from the as-deposited bar and melted to form 20 g buttons on a water-cooled copper block in a vacuum arc-furnace. Hardness tests conducted on the material before and after melting differed by only two or three Vickers Pyramid Numbers, indicating no or insignificant contamination. The buttons were hot forged, ground, and etched to sizes and shapes suitable for the rolling schedule, and vacuum annealed at 1300°F. Specimens for determination of the wire textures were reduced 91 pct in diameter to 0.027 in. in 24 steps using grooved rolls. In order for the orientation of the central region to be studied, portions of these wires were electrolytically reduced to a diameter of 0.005 in. using the procedure described by Sutcliffe and Reynolds.&apos; The sheet textures were determined on titanium cold rolled 97 pct to a thickness of 0.005 in. A reduction of approximately 10 pct per pass was used, and the rolling direction was changed 180" after each pass. Specimens used for determination of the recrystallized textures were annealed in evacuated quartz tubes at 1000°, 1300°, and 1500°F. The grain size of the 1000°F specimen was sufficiently small to give satisfactory X-ray patterns with the specimen stationary. However, it was necessary to scan the surface of the other recrystallized specimens. The microstructure of each annealed specimen was that of a recrystallized material. The diffraction rings all showed the break-up into spots typical of recrystallized structures.

Jan 1, 1954

By J. J. Pye, J. B. Drew

Mzasurements of the surface diffusion coefficients of metals have been made. Diffusion profiles for the Ag-Ag system were obtained by means of a radioactive point source and a precision auto-radiographic technique. The activation energy for silver self diffusion (=8.1 kcal per mole) is lower than that previously reported (-10 kcal per mole) on poly crystalline wire by Nickerson and Parker. The bresent data indicate an effect due to parasitic volume diffusion at temperatures above 500°C. RELATIVELY few measurements have been made of the surface self-diffusion coefficients of metals. Nickerson and arker&apos; measured the diffusion of silver over the surface of poly crystalline wires and estimated that the activation energy was 10.3 kcal per mole. Winegard and chalmers2 carried out measurements on both polycrystalline and single crystal surfaces but did not report a value of the activation energy. They found, however, that at temperatures between 250" and 400°C the diffusion coefficients were on the order of lo-&apos; sq cm per sec and that there was an acceleration of the migration of silver on the polycrystalline sample when a change of surface shape occurred. Winegard and Chalmers used an autoradiographic technique, hereafter designated ARG, and Nickerson and Parker used a surface scanning geiger counter in order to determine the diffusion profiles. More recently, Hackerman and simpson3 measured the surface self-diffusion coefficient of copper at a single temperature (750°C), and the value of the diffusivity (- 10-5 sq cm per sec) is in agreement with that given by jostein from his thermal grooving measurements. This paper reports the results of an investigation of the surface self-diffusion coefficients of silver over a large temperature range and describes the adaptation of autoradiographic (ARG) techniques for the determination of diffusion profiles obtained from a radioactive point source. EXPERIMENTAL PROCEDURE The experimental procedure is a modification of the method employed by Hackerman and simpson3 in their measurements on copper. A brief description of their technique is as follows: A radioactive needle which sinters to the surface during the diffusion an- neal serves as the source of diffusing atoms. After the diffusion run the needle is removed and the surface is scanned with a shielded counting arrangement. The diffusion profiles reported in this paper were obtained by a modification of the above procedure which employs a precision ARG technique. Previous investigations in this laboratory and elsewhere51B have shown that under carefully controlled developing conditions and by the use of calibration sources a linear relation exists between the concentration of the isotope and the photographic density for values below unity. The use of ARG under these conditions has advantages over the counter scanning method in that cumbersome shielding and requirements for great mechanical precision of the scanner are eliminated. Also the ARG gives a complete picture of the surface which is advantageous in studies of anisotropic diffusion. A recording microdensitometer having a 0.1 p wide slit was employed. At low temperatures the disturbing effects of subsurface radiations are negligible. The diffusion anneals are carried out in the cell shown in Fig. 1. The needle is formed by grinding down a 1.0 mm rod of high-purity silver until a tip of 0.2 mm radius or smaller is formed. This tip is plated withA"&apos; which becomes the source of the diffusing atoms that are detected by ARG. The needle carrier and the crystal holder, Fig. 1 are constructed of quartz and ports are provided in the holder pedestal which allow free vapor circulation ((2.0 oz) and the carrier apron fits snugly over the crystal holder cap, insuring that the needle does not move and scratch the surface. Temperatures are provided by a stabilized tubular furnace which can be quickly positioned around the cell, thus bringing the crystal up to temperature in a time that is short compared to the diffusing times. The diffusion anneals range from 2 hr for the high-temperature samples to about 25 hr for those at the lowest temperature. The possibility of vapor transport of the radioactive metal as a contributing factor in the diffusion profile was investigated in two ways. One method was to suspend the needle directly over a dummy sample, raise the temperature, for periods of time equal to the diffusion times, and then take an auto-radiograph of the surface. Negligible radioactivity appeared. In the second method a thin slot in the crystal face on one side of the source provided a "cong path" for surface diffusion. If evaporation was the primary source of surface atoms the region of radioactivity around the source would be symmetrical. This was not the case. The profile dipped abruptly at the edge of the slot but on the other side of the source the usual diffusion profile appeared.

Jan 1, 1963

By T. S. Lovering

GEOCHEMICAL prospecting extends the age-old method of searching out lodes with a gold pan and rationalizes the prospector&apos;s hunch that certain plants are associated with ore. It uses sensitive but cheap and rapid analytical methods to find the diagnostic chemical variations related to hidden mineral deposits. Exploration geologists can gain tremendous assistance from this new tool, although its optimum use is not simple. To bring out the geochemical pattern that reveals the presence of a hidden ore deposit with a minimum number of samples requires a combination of shrewdness, chemical knowledge, and exploration geology. The use of sensitive analytical methods for prospecting had its start in the 1930&apos;s in northern Europe, where Scandinavian and Russian geologists had some success in these early efforts. Very little geochemical prospecting was carried on in the United States at this time, and no sustained interest was manifest until the close of World War 11, when geochemical investigations were started by the Mineral Deposits Branch of the U. S. Geological Survey. The purpose of these investigations was to apply geochemical principles and techniques to surface exploration for mineral deposits. Both the research on analytical methods and the routine trace analyses for field investigations were at first conducted by a single group, but it later became apparent that the trace analyses could be done by men of less experience than that required for successful research on methods. For the past several years there have been two groups of chemists, and although their functions overlap, three of the chemists are chiefly concerned with research, while four to six other men make the trace analyses for field projects. The chemical investigations, as well as the field projects of the Geochemical Exploration Section, concern only those phases of the subject that are appropriate to a government organization; every effort is made to help private industry, but not to compete with it, in finding orebodies. The chief aim of the Section, therefore, is to develop new analytical techniques and publish the results promptly, to carry out field investigations of the fundamental principles of geochemical dispersion, and to field test promising- techniques under controlled conditions. Some routine geochemical exploration work is carried on in connection with DMEA loans, and in district studies where the project chief wishes geochemical information on certain areas for his report. It should be emphasized, however, that geologists of the Geochemical Exploration Section are primarily concerned with fundamental principles underlying the distribution, migration, and concentration of elements in the earth&apos;s crust. To facilitate the use of geochemical methods the USGS has published much information on its methods of analysis and has provided opportunities from time to time for qualified professional personnel to study these methods, to work in the USGS laboratory, or to attend demonstrations of the analytical techniques at the Denver Federal Center. Typical of the research carried on are the problems now being investigated: 1) Development of rapid and sensitive analytical methods suitable to the determination of traces of metals and other minor elements in various materials, such as rock, soils, plants, and water. At the present time attention is being concentrated on U, Bi, Cr, and Hg, and satisfactory rapid trace analytical methods are virtually perfected for U and Bi. Good methods are also available for: Cu, Zn, Pb, Ni, Co, As, Sb, W, Mo, Ag, Nb, Ge, V, Ti, Fe, Mn, S, and P. 2) The relation of geochemical anomalies in plant materials to the geochemical distribution of elements in soils surrounding the plant. 3) A study of the dispersion halos in transported sedimentary cover such as glacial drift and alluvium over known orebodies. 4) A study of the behavior of ore metals in the weathering cycle. 5) A study of the behavior of the ore metals during magmatic differentiation. This requires a study of the distribution of minor metals in fresh igneous rocks and their component minerals in a well established differentiation series and in adjacent country rock. 6) A study of the dispersion of metals in primary halos in the wall rock surrounding orebodies. 7) Regional and local studies of the metal content of surface and groundwater in mineralized and barren areas. Many field projects of the Mineral Deposits Branch also require the services of USGS chemists during their investigation of the geochemical environment of ore deposits. From the work that has been done certain general principles have emerged. Concentrations of an element that are above the general or background value of barren material are called positive geochemical anomalies or simply an anomaly, whereas values less than background are called negative anomalies. The anomalies most commonly investigated in geochemical prospecting are those formed at the earth&apos;s surface by agents of weathering, erosion, or surficial transportation, but more and more attention is being given to primary anomalies found

Jan 1, 1956

By R. H. King

Recently, the North American mining industry experienced a severe recession, forcing managers to take dramatic steps to cut costs and compete in the difficult international market. Some of these steps were closing mines, reducing work forces, renegotiating wage agreements, and purchasing the most productive equipment. Managers are now looking beyond these traditional avenues and are focusing more on advanced technology. In the present environment, it is essential that mine operators obtain the maximum use of capital expended on equipment. However, mine workers do not obtain maximum efficiency or productivity from equipment because the adverse and hazardous mine environment impedes human performance. Also, the efficient use of large, complex machines calls for levels of precision that many times are beyond the capability of even highly trained miners. A possible solution to this problem is a machine that mines without operators - autonomous mining machines. However, numerous problems confront researchers who are attempting to develop such equipment. Some mining tasks are not always composed of a series of cyclic motions readily performed by factory floor robots. In addition, mining takes place in the geological environment where conditions are highly variable and unpredictable. As a result, these machines must be able to sense and adapt to variations in operating tasks and environment. Considerable autonomous vehicle research has been completed, especially for defense. However, autonomous vehicle (AV) technology has not advanced to practical application yet (King, 1988) and mining research is necessary in areas: representing mining specific knowledge; •analyzing and reasoning about sensor data in the mining environment; and •discovering completely new mining methods or new approaches to existing methods that become apparent when we remove the constraints imposed by the necessity of human operators. A specific machine, the LHD, can be used to show the problems for researchers who are attempting to develop autonomous machines for mining. This author chose the LHD because it can borrow concepts developed for autonomous navigation by military programs. But considerable mining research is also required. Furthermore, studies done at the Henderson mine, in Colorado, show autonomous LHD &apos;s promise cost and safety benefits (King, 1988). At Henderson, LHDs load ore from draw points, tram to an ore pass, dump and return, or switch to another draw point. An autonomous LHD must sense vehicle position along the route, relate sensor data to stored-map information, to determine location, follow drift center lines, plan paths between dump position and initiate appropriate control commands, sense vehicle operating status and vehicle health, key on features or targets for special tasks like high speed turns, perform end of travel tasks (loading and dumping), and detect and avoid obstacles. These goals are similar to those for shuttle cars, trucks, and front-end loaders. Therefore, much of the technology is transferable. Each Henderson LHD extracts six to 40 dippers from each of a series of draw points. The LHD transports the ore to an ore pass within 55 m (180 ft) of the draw point, making the longest round trip 110 m (360 ft). The LHDs have very fast hydraulic dumping and loading systems that reduce the round trip cycle to less than one minute. Even though the LHD is capable of 500 trips per shift, the average production is 300 dippers. Man trip and lunch reduce available operating time to 6.5 hours per shift. Mucking difficulties (setting large boulders aside), operator breaks for activities like talking shop, and cleaning and smoothing roads further reduce operating time. Supervised autonomy can reduce the number of operating units by increasing operating time per shift since computer controlled machines can operate during lunch and between shifts and reduce operator errors. If dippers per shift increase from 300 to 350 (long-range goals are 500 dippers per shift), constant production requires only 10 operating LHDs and two spares. Manpower requirements drop from 24 to four by controlling five machines from one workstation. A review of technology available from the Autonomous Land Vehicle, the Advanced Ground Vehicle Technology, the Ground Surveillance Robot, and other programs, show the following differences between others work and mining industry needs: •If we focus initially on mobile haulage vehicles, we can navigate from a map. We do not need to explore. •We can modify the environment to reduce the navigation requirements. •We have a harsher environment than any of the research programs have encountered. •Our equipment must operate faster and more precisely than present AVs. •Our equipment must operate reliably over long periods of time. •We must have better onboard machine health monitoring and diagnostics. •The AV programs do not address geosensing. •The major AV programs are not cost-effective. For example, we cannot afford the computer power for robust image processing, yet. To computer control an LHD, we must replace the guidance and monitoring skills of experienced operators with sensors, computing hardware, interfaces, and several software mod¬ules. Experienced operators avoid collisions and load efficiently in piles that may contain oversize muck. Collision avoidance without an operator requires sensing all obstacles in the draw

Jan 1, 1991

By H. Sigurdson, C. H. Moore

Extensive studies have been carried out on electric furnace and blast furnace slags obtained in the winning of iron from its ores. These slags normally consist of elements of the gangue minerals present in the ores, as well as the added flux materials. In consequence, melts of CaO, MgO, Al2O3 and SiO2 can be considered as representing typical slag compositions. When a slag of this composition cools, it usually crystallizes according to predictions possible from an equilibrium diagram of these constituents, providing the melt is not undercooled to form glass. The melt is either viscous or fluid, depending upon the ratio of binary cations to silica, and crystallizes easily or forms a glass for the same reasons. If the melt is not overheated so that carbides of the metal components of the slag are formed and if the composition of the slag is so adjusted that it has a high fluidity, liquid equilibrium is attained and the slag can be held in a liquid state for extended periods of time. Upon tapping, the slag crystallizes into minerals, the type and proportion of which are determined by the melt composition. Since equilibrium is attained, the holding period is not critical. In melts containing a large increment of titanium, however, the normal slag procedures are not applicable. Titanium, as one of the atomic transition elements, is, at elevated temperatures, capable of being reduced to form metalloid compounds much more readily than the refractory oxides present in normal slags. In consequence, an oxide melt containing titanium never reaches equilibrium in a reducing environment, but continues to shift its composition until cooled. If melts of this nature are cooled and samples submitted to metal-lographic and X ray analysis the course of reaction and crystallization in this type of slag can be determined. Preparation of Slag The slags investigated fell into the system CaO-MgO-TiO2-Al2O3-SiO2 and were produced from ilmenite ores reduced by carbon in an electric furnace. Since the equilibrium series1 and the laboratory smelting of ilmenite2 are described in two of the accompanying papers, detailed description of the smelting procedure is not required here. However, certain essentials must be mentioned. Two types of melts were used to produce slags studied in this investigation. The first series of smelts made to determine proper flux addition were produced in a 4 lb Ajax induction furnace. The charge, consisting of ore with the proper flux addition, was heated in a graphite crucible until fluid, held fluid for a sufficient time period to obtain 1-5 pct FeO content, and poured. Because of the small size of the charge only the final sample of these melts could be examined. In the melts made in the 50 lb arc furnace, however, grab samples taken at 10 min. intervals between time of initial melting and final pouring were available for examination. These samples allowed a much clearer picture of the course of reaction and crystallization. amounts of ferrous oxide and reduced titanium compounds is opaque to transmitted light. Therefore, all petro-graphic studies had to be made on polished slag sections. A representative sample of slag was cut or broken, mounted in a thermosetting plastic, ground flat using 400 grit silicon carbide, the coarse scratches removed with 600 grit silicon carbide and polished on billiard cloth using levigated alumina. Rouge was avoided because of the entrainment of the red particles in pores in the slag, causing a possible confusion with some of the mineral phases. In order to prevent sample projection above the plastic surface red bakelite was used to hold the sample, and backed up with clear lucite. In this manner sample labels could be permanently retained in the mounting. The polished samples were examined on a Bausch and Lomb metallograph at magnifications of 250 X, 500 X, 1000 X and 1800 X. The instrument was equipped for examination of specimens under bright field illumination and with crossed nicols. A magenta tint plate to aid in color tone differentiation was also used. Petrology of Slags In order to determine the composition and mineral relatinos of a previously unreported system petrologically, it is essential that the starting composition, reaction temperature and final composition be known. The chemical composition of the ilmenite ore used in these smelts is given in Table 1, and the complete analysis of a typical high titanium, low iron slag is given in Table 2. In the winning of TiO2 from ilmenite by a smelting process it is necessary to produce a slag which will melt at an economically feasible temperature, remain molten as the iron is removed by reduction, be fluid enough to be readily removed from the furnace, contain a high percentage of TiO2 and a low percentage of reduced titanium com-

Jan 1, 1950

By A. J. Shaler

The subject of the mechanism of sintering has received much attention in the past few years, particularly since the beginning of the series of AIME seminars in powder metallurgy of which this paper introduces the fourth. In the first of these, F. N. Rhines1 brought together and discussed the available experimental data on the sintering of pure metallic powder, and succeeded in bringing to a sharp focus the attention of workers in this field on the established observations which a satisfactory theory must explain. Several other authors3,5,6 have, in the last few years, studied the phenomena that occur when cold metallic powders, loose or in the form of compacts, are first brought to elevated temperatures. Some workers&apos; in the field of friction have recently studied the adhesion of solid metal surfaces when they are brought into close contact. These researches have indicated that several separate mechanisms operate simultaneously, at least during the first part of the sintering process. Some of them have been called transient mechanisms4 because they are in general not absolutely necessary to sintering. Powders may be so prepared and so treated that these transient phenomena do not take place during subsequent sintering. This does not mean, of course, that their industrial and scientific importance is any less than that of the steady-state phenomena. The latter are changes that go on during sintering no matter how the powders are made or treated; they cannot be divorced from sintering. One way to analyze the process of sintering into its component parts is perhaps to distinguish between these transient and steady-state phenomena. Some of the transient phenomena have been studied in the past few years. Huttig3 has shown that, when the temperature of metallic powder is slowly raised, the following events generally occur in order: (1) physically adsorbed gases are desorbed; (2) there is an atomic rearrangement of the surface, a sort of two-dimensional "surface-reciystallization"; (3) there is a breakdown of chemically adsorbed surface compounds; (4) there is a recrystalliza-tion in the volume of the metal. All these changes are shown by Huttig and his coworkers to be completed fairly rapidly at lower temperatures than those generally used in sintering and are therefore not a part of the mechanism whereby the density of a mass of powder continues to change after long heating at an elevated temperature. But the first and third of these changes release gases in quantities which may or may not help to control the steady-state mechanisms, depending on when the voids become isolated from the outside of the compact. Among the phenomena studied by Steinberg and Wulff,8 there is the effect on sintering of residual stresses arising from the pressing operation. They found that the lateral surfaces of a green compact of iron are under a longitudinal residual tension-stress of the order of magnitude of half the yield-point for solid iron. If the outside surface is in tension, the core must be under longitudinal compression. When the compact is heated, the surface residual stress is thermally relieved first, and the compact therefore initially expands in the direction of its axis. This is a transient phenomenon, if for no other reason than the possibility of sintering unpressed powders, as demonstrated by Delisle,9 Libsch, Volterra and Wulff10 and others.1 The subject of recrystallization is dealt with further in a separate section, in view of its prominent place in sintering literature. It, too. is one of these transient phenomena. Among the steady-state parts there may be distinguished the attraction between particles and its consequences, the spheroidization of voids in the compacts, and the densification or swelling of the compact. There is considerable evidence4,7 showing that cold metallic surfaces, when brought to within a few interatomic distances of one another, are attracted to each other by forces of the order of many thousands of pounds per square inch. A calculation, discussed in greater detail in another section, shows that this force changes but slightly when the temperature of the surfaces approaches the melting point. Actual measurements of forces of adhesion of this magnitude have been made by Bradley12 on some nonmetals, but none has yet been made on cold or hot metals. This force is of sufficient magnitude to cause some plastic deformation in powder compacts, as will be shown below. A second force of steady-state nature is due to the surface tension, which probably has the same origin as the force of attraction between surfaces.164 A paper by Udin, Shaler, and Wulff1,3 gives the results of precise direct measurements of its value for solid copper. The demonstration of the tendency for the surface tension to shrink a pore was long ago given by Gibbs.17 He showed that its effect on a curved surface between two phases is equivalent to a pressure perpendicular to that

Jan 1, 1950

By J. P. Ferro, W. H. Stewart

Wet ground and dry muscovite mica continued to be the most commercially significant types of mica in the US. Canada&apos;s phlogopite mica and some US deposits of sericite mica have also contributed to the overall application of mica in a variety of industries. Mica&apos;s major end uses are paint, rubber, and construction material. Its value was about $30 million last year. The southern Appalachian Mountains weathered granitic bodies and pegmatites continued to be the primary US muscovite mica source. North Carolina production of mica as a coproduct of feldspar, kaolin, and lithium processing accounted for more than 60% of the total output. New Mexico, South Carolina, South Dakota, Georgia, and Connecticut accounted for the rest. Flake mica was also produced from mica schists in North Carolina and South Dakota. It is also being investigated in Ontario, Canada. Wet ground mica Wet ground mica was produced by four companies: KMG Minerals, Franklin Mineral Products, J.M. Huber Corp., and Concord Mica. KMG and Franklin Mineral Products accounted for more than 80% of the production. Wet ground mica is a highly delaminated platey powder used to reinforce solvent and aqueous system paints for increased weatherability, durability, and greater resistance to moisture and corrosive atmospheres. In plastics, it is an excellent filler and reinforcing agent, providing better dielectric properties, heat resistance, and added tensile and flexural strength. In the rubber industry, wet ground mica is used as a mold lubricant to manufacture molded rubber products, such as tires. It also acts as an inert filler that reduces gas permeability. Miscellaneous uses include additives to caulking compounds, foundry applications, lubricants, greases, silicone release agents, and dry powder fire extinguishers. Wet ground mica prices range from $353 to $496/t ($320 to $450 per st) fob plant. Specialty products may be higher, depending on customer requirements. Dry ground muscovite mica Dry ground mica was produced by nine companies: KMG Minerals, Unimin, US Gypsum, Mineral Industrial Commodities of America, Spartan Minerals Corp., Asheville Mica Corp., Deneen Mica Co., Pacer Corp., and J.M. Huber Corp. Dry ground mica&apos;s primary market is wallboard joint compound. Here, it is a functional extender that improves the physical properties and finishing characteristics of the mud. It is also used in various grades as a filler in asphalt products, enamels, mastics, cements, plastics, adhesives, texture paints, and plaster. Dry ground mica became popular as an additive in oil well drilling fluids, where the mica flakes platey nature helps seal the well bore, preventing circulating fluid loss. But oil&apos;s dramatic price drop and consequent curtailing of well drilling brought this once booming market to a virtual halt. Forecasters predict that this business will gradually pick up during the next few years and most current dry ground mica producers will again produce the oil well drilling material. Dry ground mica prices range from $110 to $420/t ($100 to $380 per st) fob plant. High quality sericite mica, sometimes referred to as an altered muscovite, was mainly produced by two US companies. Mineral Industrial Commodities of America and Mineral Mining Corp. have equivalent capacities of about 27 kt/a (30,000 stpy). The majority of the material produced was consumed by the joint compound industry. Minor uses are in paint and oil well drilling. The lack of ground sericite penetration into the traditional ground muscovite markets is attributed to high silica content, typically in excess of 20%, and a bulk density. Prices range from $88 to $187/t ($80 to $170 per st) fob plant. Phlogopite mica is a dark colored, magnesium bearing mica rarely found in the US. Suzorite Mica Corp., a division of Lacana Petroleum, mines a deposit in Quebec that is 80% to 90% phlogopite. The dark color has prevented the material&apos;s entry into the traditional paint markets. But the physical properties and high purity make it useful as a low-cost reinforcing filler in many plastics and several asphalt applications. Phlogopite mica is ground to several grades and may be treated with various surface coatings for use in plastics or coated with nickel for EMI/RFI shielding applications. Prices for phlogopite products range from $144 to $580/t ($104 to $580 per st) fob plant. As in recent years, production of domestic muscovite sheet - block, film, and splittings - remained insignificant. These resources are limited and uneconomic due to the high cost of hand labor required to process sheet mica in the US. Imports from India and Brazil were the primary sources of the estimated 1 kt (2.4 million lbs) valued at $2.5 million consumed by US electronic and electrical equipment manufacturers in 1986. Reserves As a feldspar, kaolin, and lithium industry coproduct, flake mica will continue to provide a large percentage of mica re- This summary of 1986 mica activity was received too late to be used in the June issue.

Jan 7, 1987

A Seminar to discuss &apos;Reactivity of Line in Steel Making&apos; was held in the auditorium of Iscor&apos;s Headquarters Building in Pretoria on the 5th August, 1971. The Institute organised this Seminar on the occasion of the visit of Dr Obst of the Rheinische Kalkstein Werke GMBH, Wuelfrath, West Germany. The latter is one of the largest lime companies in the world and Dr Obst is in charge of their research laboratory. He and his co-workers have earned a reputation for their research into reactivity of lime in steel making. This work is also of great importance to the steel industry in this country, particularly with the changes which are occurring with the introduction of basic oxygen furnaces. Some 80 persons representing most of the major steel companies attended this Seminar. Dr Obst presented his most interesting paper which he illustrated with projector slides. This paper is printed in full in this journal. This was followed by several contributions and active discussions. The first contribution was made by Dr T. B. Beeton of the Research and Development Department, Iscor. He gave a brief summary of the studies on reactive lime carried out at Iscor since 1968. Their work was particularly related to the selection and definition of the various reactivity tests and the correlation of these with plant results. He demonstrated that the DIN methods, particularly the Din (40°) test, gave the best results and reported that good agreement had been reached between Iscor, Highveld and Northern Lime on future test procedures. Mr B. Lowther of The Northern Lime Company then dealt with problems associated with the production of high reactive lime. The rotary kilns at The Northern Lime plant at Lime Acres are particularly suitable for the production of high-reactive lime, because this type of kiln lends itself to accurate control of temperatures and retention times. Several problems, however, exist. Because of the long haul distances and the deterioration of reactive lime by absorption of CO2 and water, high reactive lime is not suitable for most customers. To constantly be changing kiln conditions to produce relatively small quantities of high-reactive lime, therefore, causes problems. Highveld Steel at present is the only company requiring high-reactive lime and a satisfactory product is now being produced at a cost because kiln capacity is reduced and fuel costs are higher. Dr P. R. Jochens, of the National Institute for Metallurgy then submitted data to show how the reactivity of soft burnt reactive lime deteriorated due to long railage hauls. Detailed sampling of trucks had been carried out leaving the lime works and on receipt at Highveld and it was demonstrated that there was deterioration in the top layers of the truck. This top layer protected the balance of the lime which is still in good condition on arrival at Highveld. Exposure to atmosphere by further handling and storage into the bunkers at Highveld, however, resulted in further severe deterioration in the lime reactivity and the lime when finally fed to the Basic Oxygen Furnace is reduced to one third to one quarter of the original reactivity value. The same deterioration does not occur with normal run of kiln lime, which is burnt harder, but still has a satisfactory reactivity for most purposes. Mr A. A. Hejja of the Department of Metallurgy, Witwatersrand University then made a contribution on the metallurgical aspects of the use of lime of different reactivities in the Basic Oxygen Furnace and the effect on desulphurization efficiencies. He gave a wealth of data on actual tests which had been carried out and came up with the surprising conclusion that the deterioration of high-reactive lime due to transport did not have as serious an effect on the efficiency of desulphurization as one would have expected. The efficiency was good provided high-reactive lime in the first instance had been produced, irrespective of whether it had deteriorated thereafter. Poor efficiency was obtained only when the lime in the first instance was produced in an unreactive form. The temperature rise in the Basic Oxygen Furnace also has a definite effect. If lime is introduced when the temperature is too high, soft burnt reactive lime becomes hard burnt and unreactive before it can dissolve in the slag. Use of ferromanganese slag assists dissolution of lime. There was a great deal of discussion following these contributions. Particular interest was expres8ed in the special lime described by Dr Obst. This consists of a mixture of fine lime which is premixed with fluxes and either briquetted or pelletised. The question of injecting finely ground lime as an alternative was also discussed. Both these systems eliminate the necessity for highly reactive lime. With smaller lime particles, larger surfaces are exposed thereby ensuring more rapid dissolution in the molten slag. The discussions generally were lively and informative and all present derived benefit from learning of the experiences of others. The valuable contribution made by our German visitor, Dr Obst, was particularly appreciated. The instantaneous English-German translation service which was provided, greatly facilitated discussions.