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By H. H. Podgurski, J. C. M. Li, Y. T. Chou, F. N. Davis, R. A. Oriani

When an Fe-2 pct A1 alloy is nitrogemted at 500ºC with a gus tnixture (NH3-H2) in which the nitrogen activity has been kept Lou] enough to avoid the formation of iron nitride, a two-phase alloy is generuled which consists of AlN particles and a ferrite phase cotaining a heavy network of dislocations. The amount of nitrogen contained in such an alloy, when equilibrated with the nitrogenating atmosphere, far exceeds both that needed to satisfy the normal solution requirements of a Fe and that needed to convert all of the aluninuwi to AlN. This excess nitrogen is accounted for as being trapped on dislocations, adsorbed at the ferrite -AlN interface, and as an en-Iuznced lattice solubility in strained ferrite. This excessive uptake of nitrogen had previously been attribuled by other investigators to the formation of a nonstoichiornetric aluminum nitride. Isotope exchange experiments revealed various amomts of exchargeable N14 present in the originally nitrided samples that could not be removed by reduction with HS at 500ºC. This exchangeable nitrogen has been identified us that bound to the AlN -ferrite interface. Estimates of inter facial areas in alloys containing -3 pct by weight of A1N are as high as 10 sq m per g of alloy. ThE first1 of this series of papers described the experiments forming the basis for the elucidation of the mechanism of the formation of aluminum nitride particles within an Fe-A1 alloy. It was found that not only are dislocations necessary for the nucleation of the AlN particles but also the nitriding reaction in turn produces a dense network of dislocations in the ferrite matrix. It was also observed that nitrogen in excess of that needed for the formation of stoichiometric AIN is taken up by the alloy without the formation of an iron nitride. The present paper is an analysis of the excess nitrogen sorbed by the nitrided Fe-A1 alloy which considers the structure generated by the formation of the aluminum nitride particles. Because of the high density of dislocations, this system has proved to be quite useful in studying nitrogen-dislocation interactions. Wriedt and Darken2 have already reported such studies in a cold-worked ferritic steel. EXPERIMENTAL Nitrogen Sorption. The specimens of this alloy were cold worked (50 pct reduction) to 0.011 in. thickness and chemically cleaned in a 2:1 concentrated phosphoric acid-50 pct hydrogen peroxide solution before nitrogenation. The flowing nitrogenating atmosphere (11 pct NH3-89 pct H2) was established before the ni- trogenating temperature was reached. The same gravimetric and gas-flow equipment described in an earlier paper1 was also used in this investigation. Changes in nitrogen concentration were followed gravimetrically when establishing the isotherms. In the isotope exchange experiments, concentration changes were followed volumetrically. In some instances chemical analyses (Kjeldahl method) were used to check for material balances in both the gravimetric and volumetric procedures. Our objective was to obtain reversible nitrogen sorption isotherms for alloys equilibrated with NH3-H2 gas mixtures over a large range of nitrogen activity* and temperature. The *Defined as equal to PN H /P3/2 ,where P corresponds to partial pressure in atmospheres. Actually the nitrogen activity, aN, in the alloy equals K PNH3 /p3/2H2, where K is the equilibrium constant for the reaction NH3 = N + 3/2H2. upper limit for nitrogen activity was below that which would produce iron nitride; for reasons which will become apparent later, most of the sorption studies were made in the temperature range between 400" and 500'C. The alloy sample studied most extensively in this investigation was given a series of successive reductions (100 pct H2) and nitrogenating treatments at 500°C to attain a stabilized structure. Presumably some dislocations were lost during these treatments. Throughout most of the sorption studies, temperature was held at ± 1°C and the gas-phase composition was held to k0.2 pet NH3. Based upon the results of numerous diffusion experiments with this alloy* the 'Results to be published in a third paper of this series. times chosen for equilibration were considered more than adequate, Nitrogen Isotope Exchange. The first exchange experiments were carried out by circulating a measured quantity of H2 and NH3* (with a predetermined isotopic *The NH3 was synthesized over a synthetic ammonia iron catalyst using H2, and N2, enriched with N15. AS the NH3 was formed it was removed continuously from the gas phase by circulating the mixture through a refrigerated (78ºK) trap. When most of the nitrogen had been converted to NH3, it was distilled from the trap into a glass storage vessel. composition, N15/N14) over a nitrided specimen at 450°C in a closed glass system. Attempts to reach the exchange limit isothermally, i.e., an isotopic ratio (N15/N14) in the gas phase identical with the extractable nitrogen in the nitrided alloy, were futile. The exchange rate between the gas phase and the alloy was too slow, involving many days. To circumvent the slow exchange with the gas phase, measured amounts of N15 and N14 were introduced into the alloy by nitrogenating with an NH3-H2, mixture at 500°C containing an N15/N14 ratio of 7.76; the charged specimens were then sealed by allowing an oxide film to form over them in air, and in this final condition the specimens were subjected to 500°C in vacuum for various times during which isotope exchange was allowed to proceed within the specimen. It was estab-

Jan 1, 1970

By Peter R. Drummond

THE final casting of refined copper has been re-J- stricted for generations by the following sequence of operations: Filling the reverberatory furnace, melting, skimming, blowing or flapping, and poling. The hoped-for 24 hr cycle, producing 300 tons or more, has been taken up largely with the necessary bat time-consuming tasks of cleaning the bath, sulphur elimination, and in turn removal of excess oxygen to produce tough-pitch copper. Incidental to comparatively slow melting under combustion gases, copper oxides react with the furnace lining, and the slag so-formed must be completely recycled. The three-phase arc furnace has eliminated some of the cycle stages, and telescoped the remainder into a continuous operation. Electrical energy, supplied to graphite electrodes enclosed in high grade refractories, rapidly melts copper cathodes and sustains a stream of metal, containing approximately 0.01 pct oxygen, without contamination from fuel. The arc was struck on the first large electric furnace for melting copper in the United States on April 13, 1949. The earliest use of this type of furnace was at Copper Cliff, Ont., in 1936, and an admirable description of their installation has been published? Copper, melted in the Baltimore furnace, is used to cast billets, and the installation differs somewhat from the Canadian, as will be described. The arc furnace is a heavy-duty, three-phase furnace, holding 50 tons, the general outline of which appears on Fig. 1. The steel shell is 15 ft ID with a bottom radius of 14 ft 2 in. The roof, separate and distinct from the body, consists of a 15-ft water-cooled, cast-steel ring of the same outside diameter as the furnace. The center line of the furnace lies 9 ft 6 in. from that of the trunnions, permitting a 5" backward tilt for skimming, and a 40" maximum nose tilt forward for complete draining. Normally, the furnace overflows by displacement, and the use of the forward tilt arrangement is restricted to covering charging delays. The charging slot, 3 ft 8 in. x 5 in., lies on the north center line, the tap hole on the south, and the 30x30 in. skim door 45" to the west of the slot. The original 20-in. graphite electrodes were replaced with 14 in. in December 1949. Three conventional direct current winch drives, governed by electrical controls, position each electrode which has individual mast supports and counterweights. An independent circulation supplies cooling water for the electrode glands, the roof ring, charge slot, and the skim door frame. Arc Furnace Refractories Hearth: Fused-in monolithic bottoms had been used in copper arc furnaces, installed prior to April 1949. These consisted of thin layers of periclase, successively fused in place over preliminary brick courses. Heat was obtained from the arc, using a T-like arrangement of broken electrodes resting directly on the periclase to be fused. The operation, taking weeks to perform, was very expensive. The chemically-bonded magnesite-brick bottom, installed at Baltimore, was the first of its kind and a radical departure from previous practice. It consists of a 1 to 6-in. layer of castable refractory laid on the steel shell, modifying it to a 12 ft 2 in. bottom radius. Two courses of 9x2 % -in. fireclay straights and keys follow. The third course is made of 9-in. magnesite blocks of special shape to form circles of an inverted arch. It was started by a four piece keystone with skew-backs forming the outer course. The fourth course also started on a central keystone, or button, of four 90" segments, 12 in. diam x 13 Vz in. deep, and continued with 13%-in. blocks. Skewbacks at the shell completed the course to produce a horizontal surface for the side walls with a single course of No. 2 arch fireclay against the steel. Dry chrome-magnesite cement was brushed over each course after laying, and a 1-in. expansion space between the brick and the shell was filled with the same mixture. The total bottom thickness, excluding the castable material, was 5 in. of clay plus 22% in. of chemically-bonded magnesite. Tap Hole: A 5-in. OD and 3-in. ID silicon-carbide tube constitutes the tap hole and is set tangential to the upper course of the furnace bottom. Molten metal fills the tube when the furnace is level and filled to capacity. Side Walls: The lining, against the shell, consists of a 9x4Y2x3 in. soldier course of fireclay, using straights and No. 1 arches to turn the circles. A second soldier course of 9x4'/2x2'/2-in. fireclay was laid in a somewhat similar fashion. Three courses of 13Y2x6x3 in. and 9x6~3 in. of final magnesite, laid flat, completed the lining, using Nos. 1 and 2 keys to turn the circles. Cardboard spacers were placed between every two bricks in horizontal courses, and a thin coat of chrome-magnesite cement filled the joints between the firebrick and magnesite. A sprung-arch spanned the skim door with jambs of suitable magnesite shapes. Charge Slot: The slot is 3 ft 8 in. wide x 5 in. high. A silicon-carbide sill of special shapes has a 30" slope to allow cathodes to slide easily into the bath. The original arch was flat, and composed of Nos. 1 and 2 wedge magnesite with a 6-ft radius. It projected 5 in. over the sill, and, being a flat arch, gave an 18 15/16-in. opening between the inner edge and the metal line. The whole assembly was later raised 9 in., and the flat arch replaced with an arch, the lower edge of which maintained the 5-in, width from the outer to inner edges as shown in Fig. 2. A water-cooled, cast-copper jacket protects the steel shell behind the slot.

Jan 1, 1952

By J. J. Forbes

'"THE Federal Coal Mine Safety Act (public Law T. 552. 82nd Congress) was approved oil July 16, 1952. It incorporates, as Title I, the Coal Mine Inspectio1.1 and Investigation Act of May 7. 1941 (Public Law 49, 77th Congress), which gave Federal inspectors only the right to enter. coal mines for inspection and investigation purposes but no power to require compliance with their recommendations. Title 11 contains the enforcement provisions of the act; its purpose is to prevent major disasters in coal mines from explosions, fires. inundations. and man-trip 01. man-hoist accidents. At this point a brief account of events that preceded the enactment of the Federal Coal Mine Safety Act seems appropriate. The hazardous nature of coal mining was recognized by the Federal Govermment as long ago as 1865, when a bill to create a Federal Mining Bureau was introduced in Congress. Little was done, however, until a series of appalling coalmine disasters during the first decade of this century provoked a demand for Federal action. As a result an act of Congress established a Bureau of Mines in the Department of the Interior on July 1, 1910. The act made it clear that one of the foremost activities of the Bureau should be to improve health and safety in the mineral industries. One of the first projects selected by the small folce of engineers and technicians then employed was to determine the causes of coal-mine explosions and the means to prevent them. By investigations aftel mine disasters the fundamental causes and means of prevention were soon discovered, and the coal mining industry was informed accordingly. However, despite this knowledge and the enactment of State laws and the Federal Coal Mine Inspection and Investigation Act of 1941, mine disasters continued to occur with disheartening frequency and staggering loss of life. The devastating explosion at the Orient No. 2 mine on December 21, 1951, resulted in the death of 119 men. The Orient disaster rekindled the memory of the Centralia. Ill., disaster of March 25. 1947, which caused the death of 111 coal miners. These two tragedies ultimately brought about enactment of the Federal Coal Mine Safety Act. The act is a compromise measure. Senator Matthew M. Neely of West Virginia and Congressman Melvin Priec of Illinois introduced almost identical versions in the 82nd Congress, but they were considered too drastic. The final version was introduced by Congressman Samuel K. McConnel, Jr., of Pennsylvania, after considerable discussion and amendment in committee hearings. It was passed by the Congress and became effective when signed by the President on July 16, 1952. The act is somewhat limited in scope because it applies only to approximately 2000 coal mines in the United States and Alaska that employ regularly 15 or more individuals underground. It exempts approximately 5300 mines employing regularly fewer than 15 individuals underground and all strip mines, of which there are about 800. Moreover, it covers only conditions and practices that may lead to major disasters from explosion, fire, inundation, or man-trip or man-hoist accidents. According to Bureau records, such accidents have resulted in less than 10 pct of all the fatalities in coal mines. It is important to mention that the law is not designed to prevent the day-to-day type of accidents that have caused the remaining 90 pct or more of the fatalities, because it was the specific intention of the Congress to reserve the hazards which caused them to the jurisdiction of the coal-producing states. Many who opposed any Federal legislation that would give the Federal inspectors authority to require compliance with mine safety regulations claimed that such legislation would usurp or infringe upon States' rights. To assure that the principle of States' rights would be preserved, the act provides for joint Federal-State inspections when a state desires to cooperate in such activities. The Director of the Bureau of Mines is required by the act to cooperate with the official mine-inspection or safety agencies of the coal-producing states. The act provides further that any state desiring to cooperate in making joint inspections may submit a State plan for carrying out the purposes of this part of the act. Certain requirements are listed: these must be met by a state before the plan can be accepted. The Director of the Bureau of Mines, however, is required to approve any State plan which complies with the specified provisions. The Director may withdraw his approval and declare such a plan inoperative if he finds that the State agency is not complying with the spirit and intent of any provision of the State plan. When this paper was prepared, agreements for joint Federal-State inspections had been entered into with Wyoming and Washington. A few other states have indicated their desire to submit a State plan and negotiations toward that end are now under way. Reluctance to enter into such agreements may be due to the mine operators' knowledge that in the states that adopt a cooperative plan they are prohibited from applying to the Director of the Bureau of Mines for annulment or revision of an order issued by a Federal inspector and must appeal directly to the Federal Coal Mine Safety Board of Review for such action. Experience has proved that review by the Director as provided in the act is a less expensive and time-consuming procedure to all concerned than applying to the Board. Reluctance also may stem from the fact that joint Federal-State inspections somewhat restrict the movements of the State mine inspectors and tend to reduce the number of inspections of mines. Where a State plan is not adopted, the Federal coal mine inspector is responsible under the law to take one of two courses of action if he finds certain hazardous conditions during his inspections. The first action involves imminent danger. If a Federal inspector finds danger that a mine explosion, mine fire, mine inundation, or man-trip or man-hoist accident will occur in a mine immediately or before the imminence of such danger can be elim-

Jan 1, 1955

By Milton R. Pickus, Earl R. Parker

THE modern theories of creep¹-4 in general have been based upon the concept of generation and migration of dislocations, with the generation process normally assumed to be rate controlling. The theories are generally deficient in that they fail to take into account many factors that are known to influence creep. The influence of the state of the surface of the test specimen has been almost completely overlooked; yet the present report shows that the nature of the surface may, in certain cases, govern the creep characteristics of a specimen. In the period since Taylor" applied the concept of dislocations to a study of metals, a school of thought has developed that closely relates the plastic deformation of metals to the generation and migration of dislocations through the crystal lattice. It might be expected that the thermal energy required for the generation of a dislocation would be different from that for migration of the dislocation through the lattice. Furthermore, the activation energy for generation would be expected to vary for different parts of the solid metal. It has been predicted that dislocations would be generated most easily at external surfaces, but could also be activated at certain internal surfaces such as grain or phase boundaries. Within the body of the metal a range of values for the activation energy might be expected because of different degrees of disorder at such regions as grain boundaries, impurities, and second-phase particles. The particular value of the activation energy that was rate determining could then depend on the specific conditions of a test. If, for example, the surface atoms were by some means constrained, the generation of dislocations in the body of the metal might become the important factor. On the other hand, other conditions may favor generation at the surface. It is possible then that the creep behavior may not be completely determined by the inherent properties of the metal. Even the environment in which a test is carried out could have a significant effect. In fact it is conceivable that in order to obtain the maximum creep resistance from a given alloy, the surface atoms must be so constrained that the activation energy for generating dislocations on the surface is at least equal to that required for generation in the body of the metal. On the basis of such considerations, and in view of the limited number of publications discussing this subject, it seemed that an investigation of the influence of the state of the surface on creep might yield information of both theoretical and engineering interest. Experiments on single crystals, demonstrating a variation in the mechanical properties due to alterations in the surface layer, have been reported by several investigators.6-13 he results of these experiments have been briefly summarized;14 consequently, the earlier work will not be reviewed here. As an example of these findings the observations of Cottrell and Gibbons may be cited. They reported the critical shear stress of a lightly oxidized cadmium single crystal is greater by a factor of 2½ than a specimen with a clean surface. Materials and Methods Single crystals M in. in diam and 8 in. long were prepared from Horse Head Special zinc, melted under an atmosphere of helium in a large pyrex test tube, and drawn up into a long ½ in. diam pyrex tube by means of a vacuum pump. The cast zinc rods thus produced were cut into convenient lengths and sealed in evacuated pyrex tubes. Single crystals were grown by gradual solidification of the remelted rods. Cleaving the ends of the single crystal specimens chilled by liquid nitrogen proved a simple method for determining orientations from the exposed basal plane from the markings left on the cleaved surface that gave the slip directions with sufficient accuracy for the experimental work. The specimens chosen for the experiments were those having the angle between the basal plane and the specimen axis within the range of 15" to 65". Since zinc single crystals are quite delicate, it was necessary to devise an appropriate method of gripping the specimens in order to suspend them in the furnace and apply the load. Stainless steel collars were prepared having an inside taper, the smaller end of the taper being of such a size that the specimen could just pass through freely. The tapered hole did not extend the full length of the collar; a sufficient thickness of metal remained so that a hook could be attached to provide a means of applying the load and suspending the specimen. One of the collars was slipped over the upper end of a specimen which was supported vertically in a steel jig. The collar was then heated electrically until the end of the crystal melted and filled the collar with molten zinc. At this point the application of heat was discontinued, whereupon the molten zinc quickly solidified, due to the chilling effect of the jig. The specimen was then inverted and the second collar applied in a similar manner. The jig served several purposes: limiting the length of specimen that was melted, providing excellent alignment of the collars with respect to the specimen axis, and protecting the specimen from mechanical damage. Once the specimen was suspended in the furnace and loaded, it was desired to accomplish the surface treatment with a minimum of disturbance of the specimen. Around the specimen was a long pyrex tube, the upper portion of which was approximately 1 in. in diam, and in it was a copper coil of such a diameter to fit snugly against the tube. A specimen, approximately ½ in. in diam and 4 in. long, was suspended by means of a stainless steel rod so that it hung within the copper coil. The lower portion of the glass tube was approximately ¼ in. in diarn, and passing through it was a 5/32 in. diam stainless steel rod which hung from the lower specimen collar. This portion of the glass tube and the stainless steel rod extended through the bottom of the furnace. A T-connector, with suitable packing, was attached to the lower end of the stainless rod to provide a water-

Jan 1, 1952

By Douglas Ashmead

In his paper Mr. Braley makes no mention of the bacteriological aspects of the problem. It is now quite well established that certain bacteria play a major role in formation of acid mine waters, and it is a simple matter in the laboratory to show that under sterile conditions the rate of acid production from a pyrites suspension is only about one quarter of that obtained from a similar suspension inoculated with drainage from a mine producing an acidic pit water. Under sterile conditions the oxidation is due to direct chemical action and, from the evidence just given and from much other evidence, this increase under nonsterile conditions is due to certain bacteria. Experiments recently completed, and shortly to be published, have shown that this bacteriological oxidation can be prevented by the maintenance of pH conditions above 4. It was found that to raise this pH above 4 at the beginning of the experiments was not sufficient but that, due to the continuing chemical oxidation, alkali had to be added daily to maintain the pH conditions above 4. The amount of alkali added, however, over a fixed period, was only about one quarter of the alkaline equivalent of the acid produced when pH conditions were not controlled over an equal period. The opinion expressed by Mr. Braley that sodium hydroxide has little or no effect on the rate of oxidation of pyrites is not substantiated by the above experiments. The writer does not claim that these results show a practical solution to the problems, especially in abandoned workings, but feels that the application of an alkaline coating, such as lime wash, to exposed accessible workings might be well worth trying. S. A. Braley (author's reply)—In 1919 Powell and Parrl suggested that bacteria, or some catalytic agent, hastened the oxidation of pyritic or marcastic sulfur in coal. Carpenter and Herndon (1933)' attributed the action of Thiobacillus thiooxidans. Colmer and Hinkle (1947)3 observed an organism similar to T. thiooxidans and another organism that oxidized iron. Leathen and Braley 9rst discovered this organism in 1947 in a sample of water from the overflow of the Bradenville mine (Westmoreland County, Pennsylvania). They characterized the organism in 1954" and gave it the name Ferrobacillus ferrooxidans. Although Temple and Colmer (1951)' had suggested the name Thiobacillus ferrooxidans, since they claimed it oxidized both ferrous iron and thiosulfate, we have found that pure cultures of the organism do not oxidize thiosulfate, hence the name F. ferrooxidans. In 1955 Ashmead7 isolated an organism, similar to the one called Thiobacillus ferrooxidans by Temple and Hinkle, from acid mine water in Scotland. It is probable that this organism was F. ferrooxidans. In 1954 Bryner, Beck, Davis, and Wilsonh reported microorganisms in effluents from copper mine refuse. These organisms appeared to be similar but were not in pure culture. In view of this history of bacterial investigation of acid mine water and our own ten years of experience, we do not agree with Mr. Ashmead that bacteria play a major role in acid formation. We do not find that any of these bacteria will directly oxidize pyritic material. They do, however, augment the chemical formation of sulfuric acid by atmospheric oxidation. In two papers in 1953% eathen, Braley, and McIntyre discuss the role of bacteria in acid formation and postulate the mechanism through which they operate. Mr. Ashmead in his discussion of my paper has assumed that this work was carried on in the presence of acid mine water in which bacteria would be present. This was not the case. Strictly sterile conditions were not maintained, but the organisms present in mine drainages were definitely absent in these experiments. We believe that we have demonstrated that alkalis do not inhibit the chemical oxidation of pyritic material. This is also indicated by Mr. Ashmead's discussion in which he says that alkali must be added daily due to the continuing chemical oxidation. It is interesting to note that Mr. Ashmead finds that maintenance of pH above 4.00 decreases the activity of the bacteria. We have found also that a decrease in pH below 2.8 also inhibits its activity. Table XIII of published data'" illustrates the decrease in activity with increased acidity, although pH values are not given. These values are in comparison with uninoculated controls and show the marked increase in acidity up to 22 weeks but a decline at 29 weeks, at which time the experiment was terminated. It is probable that after a longer period only chemical oxidation would have continued. From our studiesv we have postulated that the iron oxidizing bacterium (Ferrobacillus ferrooxidans) oxidizes the ferrous iron, resulting from chemical oxidation, to ferric iron. The ferric iron then aids the atmospheric oxidation of the sulfuritic material and is itself reduced to ferrous iron, which in turn acts as food for the autotrophic bacteria. Study of the physiologic properties of F. ferrooxidans shows that its preferred pH is about 3.00 and its activity decreases with variation in either direction. It is extremely inactive above pH 4.00 and below 2.5. This inactivity above 4.00 is indicated by Mr. Ashmead's observations. These properties of F. ferrooxidans then correlate perfectly with our hypothesis. Ferrous iron is oxidized very slowly by atmospheric oxygen in highly acid sohtion and since the bacteria become inactive, acid is formed only by atmospheric oxidation. At a pH of 4.00 or above iron is more readily oxidized by atmospheric oxygen, but the bacterial activity is decreased. However, with a pH above 4.00 the ferric iron is removed from the field of activity since its soluble sulfate hy-drolyzes and precipitates the iron as ferric hydroxide or a basic sulfate. As we have shown in the paper under discussion, the alkali does not inhibit the chemical oxidation, and thus the acid formation continues. This

Jan 1, 1957

By N. J. Grant, B. C. Giessen, M. Morris

The slvuclures of alloys in the system In-Bi have been investigated after (levy vapid queuching from the mell (splat cooling) to -190°C. Tuo-phase fields could be suppressed over most of the tota1 concentvalion range; five melastable phases a1, a2, ?, ?1. and ß exisl, of which four have simple elementtike structures. For example, ß has the A5 struclure of white till. Crystal-tographic data and coordination numbers for these phases aye given, as well as addilional information on the equilibrium diagram, where a new phase In5Bi3 teas found. THE phase diagrams of binary combinations of B elements from the groups B2, B3, B4, and B5 are usually of a simple type and show considerably fewer intermediate phases than combinations of these elements with A metals, transition metals, or the B1 noble metals. Aside from phases such as the typical B3-B5 compounds with the ZnS-B3 structure type, few stoichiometric phases are known; among them are In2Bi and 1nBi.l Nonstoichiometric phases include ß (In-Sn), a1, (In-Pb), and the tin-rich ? phases in the In-Sn, Cd-Sn, and Hg-Sn systems.' The first two phases have a close relationship to indium, while the latter group are known to be structurally related to a Sn. Thus, there are generally no immediate experimental data for the correlation of the crystal structures of B metals and B metal combinations with certain monotonously varying parameters, such as the valence electron concentration (VEC), or the average atomic size. Such parameters have been recognized as structure determining, e.g., for the transition 4 Sn — y phase,2,3 where nonintegral valence electron concentrations are encountered. It should be possible to extend such correlations to other binary systems, if nonequilibrium single-phase alloys could be produced over broad valence electron concentration ranges, and if their crystal structures could be regarded as being imposed by their electronic states. It has been shown in the case of the tin-rich 7 phases3 that alloy phases with typical crystal structures can be produced in B metal alloys rapidly quenched from the liquid to -190°C by the splat cooling technique due to Duwez,3-8 and reviewed in Refs. 5 and 6. The In-Bi system was selected because it extends over a valence electron concentration range in which several metastable, nonstoichiometrir phases could be expected to occur. Further, the low melting points of the known intermediate phases, In2Bi and InBi, Fig. 1(a), indicated low binding energies and thus possibly low driving forces for the formation Of the equilibrium structures. EXPERIMENTAL TECHNIQUES AND RESULTS The preparation of the quenched foils followed the practice described in Ref. 3. Master alloys were produced by melting of indium (99.97+) and bismuth (99.99+) in evacuated Vycor capsules or in an inert gas arc furnace; quantities of 20 mg were splat-cooled onto copper and silver substrates held at - 190°C; and crystal structures and lattice parameters were determined on a GE XRD-5 diffractometer using Cu Ka, radiation at -190°C and at room temperature. The duplication of each run with both substrates permitted the elimination of overlap of the substrate diffraction pattern and that of the investigated substance. The XRD patterns were usually taken from sin2 " = 0.03 to 0.35; the substrate was used as a means of internal calibration. The fractional accuracy of the lattice parameters of metastable phases is approximately 10-3 In the following, all percentages are in atomic percent. The stable and metastable phases found after rapid quenching to -190°C are listed in Table I, together with estimated concentration ranges and crystallographic data. The investigated alloys and phases present in them are given in Table 11. Except for the region between indium with 33 and 50 pct Bi, where a revision of the equilibrium phase diagram became necessary to include a new stoichiometric phase In5Bi3, the diffraction patterns taken after heating to room temperature agreed with those expected from the phase diagram, Fig. l(a). This suggests that the new nonstoichiometric phases are not stable at room temperature, thus following the observations made for the y phases based on tin.3 Results concerning the equilibrium phases In2Bi and In5Bi3 and the revision of the equilibrium phase diagram which is made necessary by the inclusion of In5Bi3 will be treated first. The Crystal Structure of In2Bi. Makarov7 had identified In2Bi as belonging to the AlB2-C32 type ; however, in a later paper the structure was revised.8 In2Bio was found to be of the Ni21n-B8ß type,21 with a = 5.496A, c = 6.57.9A, and N = 6 atoms per unit cell.9 This crystal structure was confirmed in the present work; definite evidence for the doubling of the c axis, as compared to the A1B2-C32 type, and for the proposed order structure, was found in the powder pattern. The observed lattice parameters agree well with those of Ref. 8; those measured at -190°C are given- in Table I. The Crystal Structure of In5Bi3. A new, equilibrium, intermediate phase was identified in slow-cooled, as-cast alloys; it was identified as In5Bi3. The structure has been worked out by Giessen and Grant;" lattice parameters at -190°C are given in Table I. This phase is probably identical with "In3Bi2", produced by vapor deposition techniques, but not recognized as an equilibrium phase by Palatnik et a1.11 After completion of the present work, the existence of In5Bi3 has also been demonstrated by superconductivity measurements on In-Bi alloys.20

Jan 1, 1968

By A. L. Bailey, B. A. Landry

W. W. ANDERSON and G. E. KELLER*—We want to compliment the authors on this very thorough paper. It gives information which the coal industry has needed for some time. We hope that the additional information which the authors are collecting will he available shortly. The mixing and riffling procedure that was followed for experimental purposes is obviously not practical in routine float-and-sink testing because of the particle size degradation which would result in handling the sample so many times. It is important to obtain our tloat-and-sink fractions with a minimum amount of handling of material. A statement is made in the paper (p. 80) that "the variable most likely to affect the size of sample required to meet a given preassigned accuracy would be the state or degree of mixing of the coal." We agree that this is a large factor, but do not believe it is the most important factor. Our own opinion is that the most important single factor governing the total gross weight of sample that must be collected is the percentage of the weight of material in the smallest fraction that results from the screening and float-and-\ink operations. In other words, size of sample is governed by the total number of fractionations that must he made, and the distribution of material within the fractions. We can imagine a coal with perfect mixing, but with such a small amount of material in some float-and-sink fraction in one of the coarse sizes that a much larger sample would have to be taken than would be the case with very poorly mixed material, but with a large percentage of coarse material more evenly distributed in all float-and-sink fractions. Our own observation of many float-and-sink tests that we have run in our own organization on many types of coal is that the size of sample that must be used on fine size float and sink is governed more by the requirements for weight of material to be used for analysis in the laboratory than by weight of material necessary to obtain accurate float and sink percentage of weight values. In other words, it is our opinion that very small samples can be used for float-and-sink fractionation in the fine sizes, but that accurate analysis of the fractions will depend on a larger weight of sample being pulverized for the laboratory than is necessary to establish the float-and-sink distribution with respect to weight. A. L. BAILEY and B. A. LANDRY (authors' reply)—The authors thank Messrs. Anderson and Keller for their comments based on long experience. It is agreed that the involved mixing and riming technique used may be disadvantageous from the standpoint of degradation. Fortunately, the paper does point out that the extended riming was unrewarding in causing further mixing. Two large unknowns remain, however: (1) how much of the mixing from the presumed highly unmixed state in the bed was achieved toward the random state during blasting, loading, transportation, screening, and further transportation to the point where the gross sample was taken, and (2) how much of the mixing took place during the preparation described preceding riming. As has been pointed out by one of the authors.6 the degree of mixing has a very large effect on the size of sample required and there are still too few experimental data to show at what stage of coal handling most of the mixing occurs. The discussion states that the weight of material in a screened fraction, or in a float-and-sink fraction, is more important than the mixing factor. We do not believe that these factors are comparable in this instance inasmuch as our purpose was to give minimum sampling requirements to achieve a preassigned accuracy in the percentages of float, middlings, or sink, and nothing more. The gross sample had already been screened and no further division by screening was made or contemplated; also, it was not intended that the middlings and sink fractions would necessarily be adequate for percentage ash or other determination. In other words, the sample obtained by the method outlined is not intended for washability studies but only for preparation plant control. Further experimental work has been done, since the paper was prepared, to investigate the effect of increasingly larger top and bottom sizes on the variability of float, etc., of a double-screened coal from Western Pennsylvania. Results will be published and eventually attention is to be given to the preparation of sampling specifications. E. H. M. BADGER*—I should like the authors to explain more fully the fundamental assumptions on which their Eq 4 is based. The equation is of the form s2 = p(l - p) which is the usual expression for the (standard deviation)2 when the chance of finding a particular kind of particle in the sample is proportional to the number fraetion, p. But instead of the number fraction, the authors have used the weight fraction, WF/W. The chance of finding a particular kind of particle in the sample can only be proportional to the weight fraction, if the average ?eig?ts of all kinds of particles, that is, float, midlings, or sink, are the same. Surely a much more justifiable assumption would be that the average volumes of the particles are the same, and, if this is so, Eq 4 would not be true. This may be demonstrated as follows: Let be the weight fraction of float, middlings, or sink, dl the density of this fraction, and d2 the density of the rest of the coal. Then assuming that the average volumes of the pieces in the three classes are the same, the number fraction, p, is given by ? P = d1/l-?/d2 + ?/d1 = ?d2/d1 + ?(d2-d1) The weight fraction, w, in terms of p is given by ? = pd1/(l-p)d2 + pd1 = pd1/d2 + p(d1-d2) _____ [61

Jan 1, 1950

By Cloyd M. Smith

B. F. TiLLson*— The manifold difficulties of accurate anemometry in irregular sections of mine passageways, the irregular distributions of velocities in cross sections of the same, and the disturbing influence of the observer upon the air flow, all indicate an undesirable inaccuracy of results obtainable by any standardized method of traversing the section by an anemometer. It seems obvious that another, and simpler, method should be used to determine the volume of air flow in mine passages, namely: 1. At appropriate locations cement or calked framework rings should be installed permanently to equalize the irregularities of sectional contour and provide a place and means of attachment for a temporary cloth brattice which bears a rigid orifice. 2. The measurement of the velocity of air flow through the orifice may then be by anemometer or, preferably, by Pitot tube measurements of the differential pressures on both sides of the orifice in accordance with the standard practices available in engineering 1iterature. † The constants may be determined for various measuring positions in relation to the resulting "vena contracta." 3. The position of the person who makes the measurements is behind the brattice out of the air stream. The Pitot tube does not offer as disturbing an obstruction as the anemometer. A recording gauge may be employed to integrate fluctuations in air flow through that portion of the mine. No traverses are required because the reading may be at a single central point. An anemometer can be used with an orifice flow. The orifice will increase the air velocity at the measuring point, with correspondingly more accurate measurements where the normal air velocity through the passageway is low. Portable brattices might be devised with the cushioning rims which would seal against irregular rock surfaces where permanent rings were not available or feasible. The development by the Ventilation Committee of standard procedures and devices for the orifice measurement of the flow of mine ventilating air might be a desirable project for this coming year. C. M. Smith (author's reply)—Thank you for your discussion of my paper on underground anemometry. Your suggested method of measuring underground air flow is a novel one which might be applicable in some situations. It should be tested along with other suggested methods in any investigation of this subject. G. E. McElroy*—In spite of the adverse publicity that vane-anemometer methods of air measurement have had in the past and that contributed by the present paper, I endorse Mr. Erickovic's statement that anemometer traversing "has proved to be widely applicable, expeditious and simple" and add that available methods are accurate enough for the purposes for which they may be used. The fact that the great majority of minor mine officials assess relative changes in rates of air flow by comparison of crude vane-anemometer measurements, known to average 20 to 30 pct high, has no important bearing on this subject, because state inspection standards were based originally on such methods of air measurement. Federal inspection standards are based on actual rates of flow as determined by traversing, and interest in traversing methods is rapidly increasing. In considering traversing methods, three aspects are of major importance: (1) the absolute accuracy of calibrations; (2) the degree of interference with normal flow conditions introduced by traversing methods designed for accurate measurement by shaft-mounted instruments; and (3) the proper "method" factor to use for approximate measurements by hand-held instruments. With respect to absolute accuracy of calibrations, we have always placed reliance on calibrations made by the National Bureau of Standards, with which manufacturers' calibrations have usually agreed very closely. It is therefore particularly disturbing to find7 that calibrations made previous to June 1947 are presumably about 5 pct in error because of excessive registry caused by the thin flat plates on which anemometers were mounted for calibration. Velocities corrected for calibration have therefore averaged about 5 pct low in all probability. In this connection, it is interesting to note that an anemometer calibrated against Pitot-tube measurement by a single-point method in the Bureau of Mines experimental coal nine in 1923 indicated this same difference of about 5 pct and that the same instrument calibrated by a traversing method in a metal mine some months later indicated a difference in the same direction of about 4 pct. These results are reported by the Bureau of Mines.8 Regarding the degree of interference, or changes in velocity distribution, caused by the position of the observer's body in traversing operations, misconceptions seem to be especially prevalent, resulting in increasing advocacy of methods, such as the "clear section" method outlined in this paper, that cause just the type of interferences that they are designed to avoid. The degree of interference for any method may be gauged easily by a few experiments with a velocity-pressure gauge connected to a Pitot tube or with an indicating velocity meter such as the Velometer. In an experiment cited by McElroy and Richardson,# a decrease of 5 pct was noted at ten widths upstream from a 6-in. plank, whereas an observer's body at about the maximum practical distance of 6 ft downstream from the instrument is only about four widths away. In the Bureau of Standards paper previously mentioned, it is recommended that supports used in calibrations be at least 16 widths downstream. In practice, therefore, a downstream position of the observer is ruled out as far as accurate measurement is concerned. Operation of the anemometer by rigid shaft support from a point outside the section is seldom practicable; however, accurate results can be obtained, with the anemometer rigidly attached to a short shaft and held at arm's length, by an observer advancing across the traversed section while he faces the opposite wall and stands sideways to the current, provided that he keeps the instrument at least 3 ft away from his body at all times and traverses the entire section with it. If the traverse can be started with the observer in a side recess, the entire section can be covered in one operation. Normally, it would be covered in two half-sections. The presence of the observer's body does not, as is commonly supposed, increase the average velocity throughout the remaining part of the section. Rather, the velocities 1 to 2 ft on either side of his body are increased, but the distribution of velocities throughout the rest of the cross section remains normal, and a traverse made as stated gives a true average velocity for normal-flow conditions. Regarding the proper "method" factor for accounting for interference in the approximate methods of traversing with hand-held instruments, here again confusion prevails, for which the writer must assume some of the blame. Comparison of consecutive traverses made by shaft-held and hand-held 4-in. anemometers in field work after the tests reported by McElroy and Richardson' gave method factors

Jan 1, 1950

By E. J. Roberts

F. C. Bond (Allis-Chalmers Mfg. Corp., Milwaukee) —This paper considers comminution as a first order process, with the reduction rate depending directly upon the amount of oversize material present. The data show that other factors should be taken into account, and it is possible that in time these may be evaluated as simultaneous or consecutive reactions: Development of the theory of comminution has been retarded for many years by the assumption that surface area measurements constitute the sine qua non of the work done in crushing and grinding, and it is encouraging to note the belated growth of other ideas. In the Abstract the term "net power" should be changed to "net energy." Throughout the paper the term "hp per ton" should be changed to "hp hrs per ton", or "hp hr t." The term "Probability Theory" in the title does not seem appropriate, since it is not clear how the probability theory is used in developing the ideas in the paper. There seems to be a contradiction between the large calculated advantages of closed circuit operation and the statement following that the closed circuit test results showed no significant change in grinding behavior, when compared with the batch grind curves. Tables I and II show that between 75 pct and 50 pct solids the energy input required decreases with increasing moisture content and may indicate the advisability of grinding at higher dilutions in certain cases. The calculation of the hp-hr per ton factor indicates an input in the laboratory mill of only 7.32 gross hp per ton of balls; this casts some doubt upon the accuracy of the factor used, since the power input in commercial mills at 80 pct critical speed is customarily much higher. The tests show that within fairly wide limits the amount of ore in the laboratory mill may be varied and a product of constant fineness obtained, provided that the grinding time is varied in the same proportion. This has often been assumed, and confirmation by actual testing is of value. The Cavg corrections for differences between the plant and laboratory size distributions do not seem very satisfactory, since in many cases the plant/laboratory ratio is farther from unity after correction than before. The following equation has been derived from the data in Table VI: Relative Energy (log new ball diam in in. + 0.410) Input = --------------—--------------- from which the relative energy inputs for balls of different sizes can be calculated and compared. The relative energy input is unity for balls of 2.715 in. diam. The equation indicates that the work accomplished by a ton of grinding balls per unit of energy input is roughly proportional to the square root of the total ball surface area; provided, of course, that the balls are sufficiently large to break the material. The data in support of this statement are admittedly meager, but are fairly consistent when plotted. The relative grindability values listed in Table VI for 200 mesh multiplied by 4/5 apparently correspond approximately to the A-C grindability at 200 mesh.' It would seem that for open circuit tests comparable accuracy could be obtained much more simply by the old method' of plotting the test grind, extending the mesh grinds to the left of zero time if necessary, and determining from the plot the equivalent time required to grind from the plant feed size to the plant product size, using the average of several mesh sizes. The en- ergy input value of one time interval could be determined by tests on materials of known grinding resistance, and this multiplied by the interval required should give the desired energy input value. The relative grindabilities would be the relative time intervals required for a specified feed and product size. When the plotted mesh size lines of a homogeneous material are extended to the left beyond zero time they meet at one point at zero pct passing. The horizontal distance of this point from zero time indicates the equivalent energy input required to prepare the mill feed. The author's results show that the closed circuit grinding tests give about the same K values as open circuit tests, from which he concludes that open circuit tests are satisfactory in many cases. The value of the closed circuit test is its ability accurately to predict energy requirements in closed circuit grinding for both homogeneous and heterogeneous materials. If the material is homogeneous, the open circuit test gives satisfactory results; but if the material contains appreciable fractions of hard and soft grinding ore, the open circuit tests will not be accurate because of the accumulation of hard grinding material in the circulating load. Since in most cases it is not possible to determine a priori whether the material contains hard and soft fractions, the closed circuit tests are preferable and more reliable. B. S. Crocker (Lake Shore Mines, Ontario)—Dr. Roberts probability theory of grinding is very similar to our log pct reduced vs. log tonnage method of plotting and evaluating grinding tests at Lake Shore. However, although we both seem to start at the same point we finish with different end results. Shortly after publishing our grinding paper (referred to by Dr. Roberts) in 1939, we did pursue the subject of the "constant pct reduction in the pct +28 micron material for each constant interval of time. We ran innumerable tonnage tests on the plant ball mills, rod mills, tube mills with 11/4 and 3/4 balls, and lastly pebble mills, with tonnage variations from 180 tons per day to 950 tons per day. We found that when we plotted the log of the tonnage against the log of the pct reduced of any reliable mesh, we had a straight line up until 90 pct of the mesh is reduced. We have also tested this in our 12-in. laboratory mill with the same results. We have used this method of evaluating grinds for the past 8 years and developed the recent four stage pebble plant on this basis. By pct reduced we mean the percentage of any given mesh that is reduced in one pass through a mill at a given tonnage (or time). For example, if the feed to a rod mill is 90 pct +35 mesh and the discharge at 500 tons per day is 54 pct +35, the pct reduced is 90 — 54/90 = 40 pct. If the feed had been 80 pct +35 the discharge would have been 48 pct +35 or pct re- duced 80-48/80 = 40 pct as long as the tonnage re- mained constant at 500 tons per day. Thus we can easily correct for normal variation of mill feeds. This log — log relationship derived from the tonnage tests of all our operating mills has proved of tremendous help in checking laboratory work and in designing alternate layouts or new plants. The difference between the log — log and the semi-log plot is only shown up when the extremes in tonnages are plotted. When the relationship between the pct reduced and the tonnage was first investigated, we used semilog

Jan 1, 1952

By Paul Averitt

require both time and money. Any attempt to secure a quick answer will yield a figure that very likely cannot be substantiated, and certainly will not yield information in the detailed form now desired. It will be about 10 years before a completely new estimate can be prepared for the coal reserves of the United States at our present rate of progress, though of course, much information will be available in the interim. We would like to present such information as we obtain in the form most acceptable to the industry, and to that end we are working closely with coal resource committees of the AIME, the National Bituminous Coal Advisory Council, and the Bureau of Mines. DISCUSSION G. H. Cady*—Mr. Averitt's description of the nature of the coal resources investigations of the United States Geological Survey and the progress of this work is very timely in view of the general interest in the subject of coal resources. The Federal Survey is obviously taking the "long view" with respect to the appraisal of the coal resources as part of the preparation of the general geological map of the country and the attendant determination of the quantity of all mineral and fuel resources. It is apparently for regional surveys of the conventional type that the detailed procedures employed by the Federal Survey are applicable. Even if so it is somewhat unexpected to find the same criteria for evaluation will be applied to the country as a whole. The work on coal resources appraisal has been proceeding at various rates in a number of coal producing states for various periods. Undoubtedly since there has been no common practice in the method of appraisal the states each have worked independently and independently of the Federal Survey, and each has adopted certain practices to meet local conditions. I do not know how much consideration has been given by the U. S. Geological Survey to these practices. It seems probable that such consideration might result in important modifications of the standards which apparently have been set up by the Federal group in line with local requirements. It is the practice in some state surveys to present the facts in regard to the occurrence and distribution of coal beds and their variations in thickness down to thicknesses of 1 to 1 1/2 ft in relatively narrow stages up to maximum thickness. It is then possible for any one using such figures to compile quantitative estimates of workable coal at various minimum thickness limits. This is a desirable objective where surveys are detailed and information correspondingIy good. However, vast quantities of coal lie in beds that are relatively thin about which little information is available and the quantity of such coal might affect the final appraisal to the extent of several thousand million, depending upon the minimum limitation observed. Thus in the case of Illinois a difference of 1 ft in average thickness of coal in the coal field would make a difference of approximately 35 billion tons in the estimated quantity present. One is dealing with a very large area and small differences in thicknesses are very irnportant in the overall picture. The estimate of 10 years to accomplish the sort of detailed appraisal outlined by Mr. Averitt is viewed somewhat skeptically and certainly would require many more trained geologists than seem to be available. The requisite detailed mapping and study are very time consuming not only because of the geological field work required, which is difficult in certain seasons, but also in order to assemble data from company and personal files, and to study drill records and cores, and to sample and analyze the coal. Unless the staff of geologists could be increased in the order of several hundred per cent the prospects are that the appraisal of the type suggested will require many decades not just one. The permanent value of such appraisal, however, is not questioned and it should be expedited as rapidly as means allow. Satisfactory appraisal is not easily achieved and will vary from time to time. The main function of a geological survey is to collect and present the facts that will make possible appraisal in terms of existing conditions by individuals, particularly engineers, qualified by training and expcrience for making such estimates. Thickness and character of the coal bed, nature of the overlying and underlying strata, position, thickness, and character of "partings," relationship to other coal beds, and a variety of other factors must be assembled by the geologist in order to present a complete picture of the geological setting. The economic and engineering factors are in general outside the field of geological experience. The geologist should see to it that the pertinent geological facts are available; much more information than simply thickness and depth of the coal beds is essential to meet all the requirements of appraisal in the future as well as at present. With respect to the present demand by engineers and coal mine operators and others for a rapid re-appraisal of the available coal resources, in the light of existing practices and those in immediate prospect, the Federal procedure of systematic mapping and appraisal is scarcely in line. It will be necessary for those particularly interested in having such an estimate made to establish standards with respect to the various factors involved so that the scope of the geological work can be definitely restricted for the various coal fields. There is little use in setting up general standards since no such standards will be applicable to all fields. Thus it is folly to establish a standard minimum thickness to be generally applicable which is one-half to one-third the minimum minable thickness in some fields. The greatest difficulty in appraisal of the presently actually workable coal bed lies in the variation in the importance placed on roof, floor, structural conditions, bedded impurities. sulphur content and other factors that affect judgment relative to the value of a coal bed. It is doubtful whether any appraisal will have universal approval in view of the variable factors involved. For this reason if none other, the facts should be carefully and painstakingly assembled after the manner of a systematic geological survey such as that advocated by Mr. Averitt. Only upon the basis of sucb information can reliable estimates be made in line with selected conditions. Paul AveRitt (author's repty)— In this paper I discussed, but did not define as such, two major phases of our activities on coal, which are closely related, and which normally we in the U. S. Geological Survey see as an integrated whole. It is understandable, therefore, that Dr. Cady should assume that my remarks applied only to the current work on coal reserves with which I am directly concerned. In view of Dr. Cady's discussion, however, I think a brief summary of the two phases of U. S. Geological Survey activities on coal will be appropriate. The long-term function of the U. S. Geological Survey on coal investigations is to make detailed geologic studies, involving the preparation of detailed geologic maps showing the outcrops and correlations of coal beds, direction and dip of coal-bearing rocks, nature and thickness of overburden, location of faults and fold axes, and similar features. Such studies also yield many measured sections of coal and associated strata, which show variations in the thickness of the beds, and the intervals between coal beds. Taken as a whole, detailed geologic studies provide much useful information for mine operators and engineers. As Dr. Cady points out, this type of geologic work is arduous and time consuming, but it provides the basic data that are necessary both in the development of our coal fields, and also in the preparation of coal-reserve estimates, which are discussed in the next paragraph. The U. S.

Jan 1, 1950

By Asheim, Harald A.

Parameter estimates affect both the optimum field development decisions and the economic result. A general methodology for assessment of the effects of parameter uncertainty in offshore development decision situations is proposed, and used to analyze aspects of the initial development situation for the Statfjord field.

Jan 1, 1982

By P. J. Clifford, A. B. Brown

The development of computerised mine planning has been hindered by the cost of processing the large data bases associated with mineral deposits and the inherent inconvenience of performing an interactive process on a time-shared computer. A dedicated computer is ideally suited for the purpose but reasonably priced micro/mini-computers have not had the capacity to store the volume of data or the processing capability required. Hardware development has cleared the way for the development of truly interactive mine planning software, covering the complete range of activities associated with mine planning, from geological assessment to short-term scheduling of operations, incorporating them into a single system using a common database. The integrated system, comprising software and dedicated hardware, adds previously unattainable speed, flexibility and clarity to the planning process because it has been tailored exclusively for that purpose.

Jan 1, 1983

By Edwin B. Wilson, Charles F. Cockrell

The application of evolutionary operation is proposed for the cleaning of coal by the heavy media process. Analysis of the data given suggests some benefits that may not be immediately appreciated unless evolutionary operation is applied. For example, while it may seem obvious to increase specific gravity when the ash and sulfur content of the clean coal is lower than that allowed, evolutionary operation may show that an increase in feed rate or an increase in both feed rate and specific gravity would be a more profitable alternative. In addition, by using evolutionary operation, up-to-date knowledge of the effect of change of feed rate or specific gravity on yield and product quality is available. This information facilitates confidence about decisions made in regard to the day-to-day operation of these units. A number of other applications in the minerals industry are proposed. An application of evolutionary operation in a related field is also given where a large number of variables were required to be optimized.

Jan 1, 1969

By Roy R. Weil, John A. Dziubek

This paper discusses the use of computer modeling, data manipulation, and mapping of three multi-seam coal properties in the Appalachian coal basin. Each case history represents a slightly different application of computer usage for estimation of reserves and planning surface mine sequencing. This practical use of computers demonstrates the suitability of data processing for normal evaluation of coal properties.

Jan 1, 1983

From the historical account of the coal industry set forth in the preceding pages the reader will have learned that coal is extremely widely spread throughout the United States, and in most places it has been easily found, that it has been remarkably easy to develop, and, where the deposits were available to streams on which it could be transported to markets, it was opened almost as soon as the country was settled. Such were the mines along the James, Susquehanna, Monongahela, Ohio, Kentucky, Cumberland and Big Muddy Rivers. In other localities where such easy means of transport were not convenient, the early production was confined to local use, or to places to which it could be hauled by wagons, but everywhere small mines were established almost as soon as the country was settled, and these increased both in number and size as the available markets grew. These early mines were nearly always opened by local people, and the industry was so far-flung that its growth attracted little or no attention excepting when a labor disturbance or a breakdown in transportation occurred. Had it been concentrated in a few places as most metal industries were, or as the petroleum industry was for many years after it started, it is probable that much better records of its progress would have been kept. When the canal era began, in the eighteen twenties, coal was at first not considered as a valuable source of freight revenue, and much surprise was expressed that the receipts of the Schuylkill Canal were very largely from coal after the first few years, as was the case of the Union Canal though not to such an extent. Even the early railroads to the coal fields did not realize the extent to which the products of mines would figure in their revenues, and it was many years before railroads were built practically solely for prospective coal traffic; indeed this did not happen until some years after the Civil War. Map 12 shows the extent of the canal system in the United States in its relation to the coal fields at the time of its maximum development. It will be seen that only the coal areas in Pennsylvania, Ohio, and to a very small extent in Indiana were able to ship by canal at all, and that only a very small portion of such fields could do so. Looking backward, it is hard to understand the great hopes entertained of canals by their pro-

Jan 1, 1942

By M. Kashiwada

The zinc leach residues are introduced into waelz kiln to fume volatile metals and before the end of 1967, the waelz-fume containing zinc, lead, cadmium and indium was directly recycled back to the leaching process for the production of electrolytic slab zinc. However, since 1968, we have been recovering with success those valuable metals contained in the fume separately instead of direct leaching. According to the improved process, the waelz-fume, after being mixed with chloridizing agents, is fed into the rotary kiln to be roasted, wherein most of lead and cadmium content is collected as fume and most of zinc remains in the calcine. Zinc calcine thus obtained, containing 70 to 75 per cent zinc and, less than 1.0 per cent lead is sent to zinc plant for further recovery and the fume containing 40 to 50 per cent lead and 3 to 6 per cent cadmium is further processed to recover lead, cadmium and indium. The advantage of this process can be secured to the utmost when it is combined with the following subsequent treatments. (a) Blast furnace smelting of non-magnetic waelz slag to recover copper, silver and gold as matte. (b) Gallium recovering process from magnetic waelz slag. (c) Soda process of lead bearing cake produced after removing cadmium and indium from the said fume. Thus the extractions of zinc, lead and cadmium from leach residues account for 87 per cent, 72 per cent and 85 per cent respectively.

Jan 1, 1970

By P. C. Upadhyay, J. G. Singh

Reissner's thick plate bending theory has been recast for rocks which exhibit the property of double elasticity: different Young's modulus and Poisson's ratio in compression and tension. The governing differential equation and other relevent expressions have been presented in a form compatible to the applied mechanics results so that they can readily be used. Effect of the ratio of Young's modulii in compression and tension, (Ec/Et), on the maximum plate deflection and the maximum bending moments has been evaluated over a wide range. It has been found that the inclusion of bi-modulus property through a parameter (ß = Ec/Et), changes the sheer deformation of the plate to such an extent that the bi-modular effect of rock cannot be overlooked in the thick plate analysis, especially when ß is expected to be greater than 2.

Jan 1, 1982

By Michael Tenenbaum

THE importance of semikilled steel as a high tonnage grade has long been recognized. The increasing severity of the applications for which semikilled steel is used makes it desirable to obtain further information regarding the features of open hearth practice and ingot structure that can affect steel performance in subsequent rolling and fabricating operations. Accordingly, an investigation is being carried out studying the structure of various types of semikilled steel ingots. This paper reviews some of the observations made and information gathered in this investigation. By definition, semikilled steel may be considered to include all nonrimming steels in which the natural shrinkage occurring during ingot solidification is offset to an appreciable degree by gas formation. Accordingly, this type of steel includes the entire range intermediate between rimmed and fully killed steel. The resultant ingot structures range from those of steels that have been capped to suppress a weak rimming action to those steels that have been killed in the furnace or ladle to the extent that no mold deoxidizer is required. A survey reported to the 1946 AIME Open Hearth Conference' reviewed the practices being used at 28 major plants for the deoxidation of semikilled steel. It was found that most plants divide the deoxidizing additions between the ladle and the mold many of the plants adding only minor amounts of deoxidizer to the bath or none at all. Silicon, aluminum and titanium were the only elements used for ladle and mold deoxidation. Mold deoxidizers were added as capping additions or as uniformly fed additions. Several practices were reported in which the steel was deoxidized in the ladle to the extent that little or no mold deoxidation was needed. No reference was made to grades capped by using special mold designs or special closures over the top of molds. Accordingly, specific examples of this type of practice will not be considered in this paper. It appeared from this survey of deoxidation practice that semikilled steel can be divided into three types, namely, capped semikilled, in which aluminum is added near the end of pour or after shut-off, intermediate semikilled steels, in which aluminum is fed uniformly throughout the pour, and semikilled steels requiring no mold deoxidation. In an attempt to cover the ingot structures obtained by the range of practices reported in the preceding survey, three series of experimental heats were made using increasing amounts of aluminum, silicon and titanium for ladle deoxidizers. The ladle deoxidizers were added as stick aluminum, 50 pct ferrosilicon and 20 pct ferrotitanium (4 pct carbon), respectively. Two methods of studying ingot structures were used. Ingots were removed from production, burned longitudinally and machined to allow observation of the central pipe and coarse blowholes. During the machining operation much of the finely porous subsurface structure was obliterated. Accordingly, observations were also made

Jan 1, 1947

By John Cogan

INTRODUCTION Lake tapping is a method of blasting an intake into a body of water from below the natural water surface without first lowering that surface or installing a protective cofferdam around the tap hole. Lake taps are done by first excavating a tunnel almost to the water/rock contact and then blasting out the final protective rock plug at one time to allow water to suddenly inflow into the tunnel from the lake. This procedure can be done "dry" or "wet." In the dry method the tunnel is empty of water before the final plug blast while in the wet method it is partially filled with water. This paper describes a wet method lake tap as performed at Alaska Power Authority's Tyee Lake Hydroeletric Project in southeast Alaska. The Tyee Lake project is located approximately 40 miles southeast of Wrangell, Alaska, in a glacially formed valley called the Bradfield Canal. The project consists of a powerhouse located at sea level on the canal, a system of tunnels and shafts connecting the powerhouse to Tyee Lake at elevation 373 m (1225 ft), and an 129 km (80-mile) electrical transmission line leading from the powerhouse to Wrangell and then Petersburg, Alaska. Tyee Lake itself is a natural lake gouged out of the rock by glacial ice during the last ice age. The objective of the project was to tap into the lake and use the water as a source for power production. The techniques used for lake taps have been developed in Norway where conditions are favorable for their use in hydroelectric proj¬ects. Over 100 have been performed. In the United States the Tyee

Jan 1, 1984

By W. R. Crane

THE support of roof in mines is dependent largely on the character of the top rock and its occurrence. The formations overlying the orebed in the Birmingham district are sandstone and slate. The sandstone occurs in beds of sufficient thickness to constitute important elements in the system of support; the slate is relatively strong but is thinly stratified. The alternating beds of sandstone and slate furnish an excellent combination in that the former are strong and the latter are impervious; however, their continuity is broken by jointing or slip planes. The thickness, and consequent weight of the overlying formations, is important in so far as the size and arrangement of pillars are concerned but may be relatively unimportant, temporarily at least, as a factor in the support of the roof. The position of the respective beds of sandstone and slate, their relative thickness, and their inherent weakness because of the occurrence of faults and folds, may exert a predominating influence on their support in place, which is affected only in part by the weight of the cover. The only mine in this district that operates under a cover of approximately 2000 ft. (1900 ft.) is the Shannon so-called "twin slope," situated 14,000 ft. southeast of the portal of the No. 7 mine of the Tennessee Coal, Iron, & Railroad Co. The conditions in this mine indicate what may be expected in other mines when the same depth is attained; in fact the effect of pressure is shown in several mines that have not reached two-thirds of that depth. It is evident then that conditions affecting the support of roof will not improve but rather will become more difficult with the extension of the mines into the valley under a constantly increasing weight of cover. Further, the difficulties will probably be augmented by the occurrence of faults and other disturbed ground, the presence and position of which at present are largely unknown.

Jan 9, 1924