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By Thomas T., Read

ALTHOUGH the first colonists in the area that is now the A United States, whether Spanish, French or English in nationality, were usually keenly interested in the possibilities of mineral wealth, it is a curious and interesting fact that none of them happened upon the mineral deposits that eventually were to make this the greatest mineral-producing country; up to 1800 it would probably have been rated as rather poor in minerals. Spanish explorers traversed the southeastern United States with- , out finding the Appalachian gold deposits that aroused so much interest in the early nineteenth century; Coronado sought in vain for the seven golden cities of Cibola; and the important gold \ deposits of California were not discovered until the political sovereignty of the region had passed from Mexico, which had held it for three centuries. The French also held Canada for nearly two centuries (and the British for another century) without finding the mineral deposits in Ontario and Quebec that now are so important. Though they knew of the copper of the Keweenaw peninsula, the lead ores. of Wisconsin and Missouri, and the coal of Illinois in the seventeenth century, these deposits were all too remote for exploitation. In the area held by Great Britain, iron smelting was planned at Jamestown as early as 1619 and was actually begun at Lynn, Mass., in 1643. The industry grew slowly, at first, but had reached considerable importance before the Revolution. Commercial bituminous coal mining began about 1750 in Virginia but anthracite mining in Pennsylvania was not started until 1800. As a result there was naturally little interest in the technical aspects of mineral prodnction here before 1800 by any except the few people who were actually engaged in it. Students of natural science were much interested in minerals, however. The activities of those who rank as geologists are so thoroughly covered in

Jan 1, 1941

The amount of coal used in the Colonies before the Revolutionary War was very small, and few records of it can be found, these all being in the Public Records Office in London. The exports and imports for the years 1768 to 1772, inclusive, are shown in Table 3, which apparently do not include the early shipments from the Virginia field. For the three years 1769 to 1771, the imports from Great Britain-their only source-totalled 17,943 tons, or an average of 5,981 tons per year. If the shipments from Richmond for the same period were available, this average would probably not exceed 6,800 tons per year. In Table 4 the shipments of coal from the James River ports for the period 1758 to 1769 are shown. A few of them are re-exports of coal brought from England, but the great majority are undoubtedly of coal produced in Virginia, and the movement was partly to West Indian ports but mainly to Atlantic ports from Baltimore to Portsmouth, N. H. (Piscataway). In Table 5 the data in Tables 3 and 4 are combined, the exports being deducted, and this summarizes all the available data on the amount of coal consumed in the Colonies before the separation from England. As has already been stated there has been a tremendous amount of work done in the Richmond field over a long period of time, many mines were opened and in some cases operated for years, and practically no accurate data remain or have been published to show what was done. There are undoubtedly manuscripts still remaining that will contribute to much of the knowledge of this field, and it is well worth the while of the proper state authorities to collect all the information that can be found anywhere, to enable the future development of these coal deposits to benefit by the record of what was done in the past. Table 22 shows all the methods of shipment from this field, and all of the data available about tonnages moved over each of them. The shipments by the Chesterfield and Clover Hill Railroads are largely duplicated in some of the other columns, but allowance has been made for this in the estimated totals. In addition to the difficulties experienced with similar data for the other states, these Virginia records have one peculiarly their own; it was the custom to allow 5 pecks to the bushel of coal at the pit and only 4 pecks at the docks or at vessel side. In most of the figures it is not stated to which point they apply, and a considerable error may exist due to this cause, although certainly it does not apply to all of them. The notes show the sources of the figures used, and while it is probable that the totals are lower than the actual amounts produced, the data given are the most complete that can now be assembled.

Jan 1, 1942

By J. F. Shirley, H. K. Burke

In January, 1964, three years of laboratory and pilot plant test work reached its culmination. The San Manuel Div. of the Magma Copper Co. switched its molybdenite extraction circuit from the sodium hypochlorite-ferrocyanide process to the new process described in this paper. The San Manuel Div. Concentrator of the Magma Copper Co. is located near San Manuel, Ariz., approximately 45 miles northeast of Tucson. The concentrator treats approximately 35,000 tons of ore per day containing about 0.85% copper and 0.025% molybdenite. The San Manuel ore body is a disseminated copper deposit in quartz monzonite and quartz monzonite porphyry. The primary copper mineral is chalcopyrite; minor amounts of chalcocite, bornite and covellite are also present in varying degrees. Pyrite makes up somewhat more than half the sulfide content of the rock. The molybdenite is nearly always found on fracture faces or associated with quartz veinlets.' On January 9, 1964, the sodium hypochlorite-ferrocyanide process that had been used for molybdenite recovery was replaced by the following process: 1) A conditioning step with hydrogen peroxide, sulfuric acid, sodium cyanide and zinc sulfate. 2) A two stage rougher flotation step with stove oil and sodium ferrocyanide. 3) One cleaning step with sodium ferrocyanide. 4) One cleaning step with sodium hypochlorite and potassium ferricyanide. 5) Four cleaning steps with an anti-foam Exfoam 636 and potassium ferricyanide. COPPER CONCENTRATOR Since molybdenite recovery is closely allied with previous steps in the operation, the copper concentrator will be briefly described first. The ore is ground to approximately 5% plus 65-mesh and 65% minus 200-mesh in a two-step grinding circuit consisting of rod mills in open circuit and ball mills in closed circuit with rake classifiers. A straight forward rougher circuit producing only one concentrate is used. The rougher concentrate is combined with the discharge from the regrind ball mills and is then cycloned, with the overflow going directly to the cleaner circuit and the underflow going to the regrind mills. The feed to the cleaner circuit is approximately 90% to 94% minus 325 mesh. The cleaner circuit consists of two stages in which the tailing from the final cleaner stage is returned to the primary cleaner and the primary cleaner tailing is returned to the rod mill discharge at the head of the circuit. Approximately 0.013 lb of the allyl ester of amyl xanthate (S-3302) per ton of ore is added to the rod mills and approximately 0.006 lb of sodium isopropyl xanthate (2-11) per ton of ore is stage added down the rougher banks. The frother used is methyl iso-butyl carbinol (MIBC) and it is added at various points in the grinding and the rougher flotation circuits. Lime is added to the ball mills and/or to the rod mills to maintain the desired pH in the rougher flotation. The rougher flotation has been operated at various pH levels between 10.5 and 11.5. Fuel oil is added in the regrind mills at a loading of approximately 0.01 lb per ton of ore. Factors Which Directly Affect Molybdenite Recovery: It has been found at San Manuel that, to obtain maximum molybdenite recovery, an oily type flotation collector must be added in the grinding circuit. When molybdenite is crystallized in the characteristic platy form, which appears to be the case almost all of the time at San Manuel, its surface has a greater tendency to adhere to oils than water. 2 Of the several oily type collectors available on the market, the allyl ester of amyl xanthate seems to be the best molybdenite promoter. The floatability of San Manuel's molybdenite does not appear to be affected, at least up to a pH of 11.5, by lime. It has been reported that at several plant operations high lime alkaline circuits have adversely affected the molybdenite recovery.3,4 It has been found through microscopic studies that the bulk of the chalcopyrite will float before the molybdenite in both the rougher and cleaner circuits. It is necessary to add fuel oil to the regrind mills to enhance the

Jan 1, 1965

By J. E. McConnell

No. 1 Electric Furnace and Billet Casting Shop, located in Plant No. 4 adjacent to our new 12 inch merchant bar mill, represents Inland Steel Company's first venture into both electric furnace steelmaking and continuous casting. Construction of the shop began in September, 1968 with excavation and foundation work, followed by awarding of general construction and major equipment contracts from December, 1968 to April, 1969. The first heat was tapped from No. 90 furnace on December 18, 1970, and the first cast attempted on December 29, 1970. The rated shop capacity is 500,000 tons per year of semi-finished billets and blooms, and was justified on the basis of approximately 85 percent of the product mix falling into the special bar quality category of steel grades. Auxiliary equipment has been designed into the shop to pursue free machining steels, including leaded and tellurized grades. The start-up and operation of the shop for the first year was essentially based on a two-part philosophy: 1. Initially operate with the "open stream" casting technique to accomplish machine shakedown and operator familiarization and training. 2. Operate with special bar quality, or "submerged casting", practices, make necessary refinements and gain customer acceptance

Jan 1, 1972

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 Hugh P. Nicholson

IN view of the many articles that have been written on the proper classification of mining systems without so far having produced any particular tangible results it might seem that the subject is hopeless of solution. I do not believe that is so, and my intention in contributing the following outline is to present a slightly different viewpoint. Too much attention is often focused on unnecessary details and not enough on general principles. Too much attention and emphasis is often given to the system employed at one particular mine or in one particular district. In addition, names that are entirely misleading are applied to mining systems, and their adoption by the coming generation perpetuates a condition that needs clarification badly. The following classification is the result of several years of practical experience in mining and also several years' practice in teaching the same subject. It therefore represents a broad picture of the entire field of mining systems from the viewpoint of the miner as well as the teacher. It seems that the difficulty in the past has been in reconciling these two viewpoints. The need of such a reconciliation requires no explanation.

Jan 1, 1934

By Anton Sawatzky, Erich Vogt

By means of mathematical solutions to the appropriate diffusion equations, we describe the kinetics of the thermal diffusion of hydrogen in Zircaloy-2 for the various temperatures and concentrations encountered in a heavy water moderated reactor. When the hydrogen concentration is below terminal solid solubility only the a Phase is present. Redistributions are then described in terms of the characteristic functions of the difhsion equation. For higher concentrations both the a and 6 phases are present. We assume the two phases to be always in equilibrium. For moderately small hydrogen concentrations exact solutions of the two-phase equation approach the the approximate solutions derived by Sawatzky and for all concentrations the exact solutions exhibit the qualitative features of his result: the two-phase concentration increases with time, everywhere; in the absence of a hydrogen current at the hot end of the sample an a-phase region always exists there; the interface of the a + 6 , a-phase boundary moves toward the cold md of the sample and the hydrogen concentration is discontinuous at the interface. Simultaneous solulions of the a and a + 6 hydrogen distributims and of the concomitant interface motion are obtained and compared to the observations of Sawatzky and Markowitz. The kinetics of the hydrogen diffusion process are shown to lead to m apparent heat of transport of the a phase which is lower than the actual value (even for samples with long anneals) thus resolving at least partially the disparity between experimental measurements of this quantity. A number of recent papers1"4 have reported measurements on the diffusion and redistribution of hydrogen in Zircaloy-2 under temperature and concentration gradients. These studies were instigated by problems arising from increasing use of Zircaloy-2 as a fuel element cladding material in pressurized-water power-reactors. The Zircaloy-2 picks up hydrogen during the operation of the reactor: the consequent precipitation of zirconium hydride in the Zircaloy-2 has pronounced effects on its mechanical properties. The Purpose of the present Paper is to describe the kinetics of the thermal diffusion of hydrogen in Zircaloy-2 for the various hydrogen concentrations and temperatures likely to be encountered in reactors. When the hydrogen in the Zircaloy-2 is entirely in the solid solution phase (a phase), the differential equation for thermal diffusion is well known and the redistribution can be described by standard mathematical methods some of which are given in Section 11 below. The treatment of the a + 6 (hydride) region differs from the earlier treatment of sawatzky3 by taking into account several modifications of the two-phase diffusion equations as suggested to us by Markowitz4 and Kidson.5 Even with the modifications the results obtained in our more complete treatment are substantially the same as those found by Sawatzky. As we show, the difference between the results of sawatzky3 and kIarkowitz4 is largely due to the difference in the geometry of their experiments. In the next section we derive the modified two-phase equation for arbitrary geometry. The exact solution of this equation is given in Section III for linear and cylindrical geometry. It is shown that Sawatzky's approximate solution is quite accurate for almost all the temperatures and hydrogen concentrations which are actually encountered. In Sawatzky's approximate theory the hydrogen concentration everywhere in the two-phase region increases continuously. As his paper pointed out, this result, together with hydrogen conservation, implies that in a sample with no hydrogen current flowing into the hot end a two-phase region is always accompanied by a single-phase region at the hot end of the sample. The net hydrogen gain in the two-phase region is supplied by the decrease of hydrogen in the single-phase region and by the movement of the (a, a + 6) boundary toward the cold end of the sample. Hydrogen conservation at the (a, a + 6) boundary leads to a discontinuity in the concentration and its derivative there. In Section IV it is shown how these qualitative features of the thermal diffusion kinetics arise from the simultaneous solution of the a-phase differential equation and a + 6 phase equation. Methods derived by us previously for the solution of the redistribution in both phases and the accompanying motion of the boundary are used to obtain approximate solutions valid for large times. The solutions account well for the observed redistributions of Sawatzky and Markowitz. It is shown how the continuing motion of the (a, a + 6) boundary leads to measured values of the heat of transport lower than the actual value, even for specimens with relatively long anneals.

Jan 1, 1963

By R. J. Grace

Primarily, this report presents the results of direct efficiency tests which were conducted on the 200-hp Coal-Pak, hot water generator, No. 3 unit, at the Otterbein College heating plant in Westerville, Ohio. Also included is a comparison of fuel consumption of the old plant with that of the new plant, a comparison of efficiencies obtained by the indirect method on a 300-hp Coal-Pak, and the cost of gas vs coal to show the savings effected by the latter. Tests were conducted with ten different bituminous coals selected by Mr. S. A. Frye, business manager at Otterbein College. The coals, of ten different suppliers, from four different states and eight different seams, represented competitive products economically available to the heating plant. The tests were conducted by the usual plant personnel which consisted of the plant superintendent and two assistants. No outside supervision of any kind was given during the tests. The tests were made from Dec. 13, 1960, through Mar. 4, 1961. During this period, 300.03 tons of coal were burned in the No. 3 unit. This tonnage represented 25.5 pct of the total coal requirements for the 1960-61 season. Operation during the period was typical for the comfort heat load which the plant supplies. Of the 78.12 operating days, the boiler was on bank or holdfire 14.31 days, or 18.32 pct of the time. The length of individual bank periods varied from 2 hr to three days. Run periods varied from about 2 hr to 7.26 days. DESCRIPTION OF PLANT AND HEATING EQUIPMENT The new Otterbein College heating plant, completed in the summer of 1957, is located in a 50x50 ft boiler room on the lower level of a combination maintenance building and heating plant. The building is situated away from the main college complex on a hillside, which allows for on-grade delivery of coal to the three bunkers. The bunkers are located directly above and connected with the individual stoker hoppers. Fig. 1 shows an in-plant view of the Coal-Pak installations. The No. 3, hot water, unit is at the extreme right. The heating plant consists of three 200-hp Coal-Paks. Unit No. 1 is operated at 4 psi and carries the steam heating load of 14,000 sq ft EDR (steam) installed in two buildings. Provision has been made to convert these buildings to hot water heating at some future date. Since the unit under maximum load conditions operates only at about 50 pct steaming capacity, a heat exchanger was installed to convert steam to hot water. Thus, during light loads in mild weather the unit can be employed more efficiently to supply both steam and hot water and eliminate the need for operating either of the two hot water units. Units No. 2 and No. 3 serve a hot water heating circuit for a large group of 21 buildings of the College. These boilers deliver hot water at temperatures up to 200°F to the various buildings through distribution lines in tunnels. The present calculated maximum heat loss for the College is 15.195 million Btu per hr. The Coal-Paks have individual design capacities of 6.700 million Btu per hr. Table I lists the design date for the Coal-Paks. The No. 3 unit was chosen as the test unit, primarily because of its bunker location at the far end of the delivery ramp. Because of this location any excess coal above bunker capacity could be conveniently stored and isolated from the adjoining bunkers. Each bunker has a capacity of 30 to 35 tons of coal. The No. 2 and No. 3 hot water generators are equipped with automatic controls which maintain the

Jan 1, 1962

By I. H. Warren, E. Devuyst

A review of the physical and chemical aspects of the direct leaching of metal oxides has been given and com- pared to recent data of the authors and co-workers. The physical aspects, including the anisotropy of the dissolution, the hydration and the zero point of charge of some metal oxides are discussed first and then the kinetics of the dissolution of metal oxides are considered. A general mechanism of the leaching of oxides which describes most of the observed results is proposed.

Jan 1, 1973

By Samuel S. M. Chan

The magnetic susceptibilities of some ores and the major rock formations of the Precambrian Belt Supergroup in the Coeur d'Alene mining district were determined both in the laboratory by the use of the Gouy method and in the field with a Bison susceptibility unit. Trace element analysis of the same group of rock samples was also made using the atomic adsorption technique. Results of this preliminary study indicate that some correlation exists between the magnetic susceptibilities and the various rock formations. This may be useful for regional mapping and mineral exploration.

Jan 1, 1974

By H. W. Knott, M. H. Mueller

The crystal structures of Ti2Cu, Ti2Ni, Ti4Ni2O, and Ti4Cu20 have been determined using powder specimens examined by X-ray and neutron diffraction. Lattice constants have been determined for all four phases using X-ray powder diffraction films. Atom positional parameters of all four phases have been determined from observed neutron intensities. X-ray diffraction calculated intensity data have been presented also for the phase Ti2Cu to point out the particular suitability of neutron diffraction in this case. Interatomic distances have been determined using the positional parameters obtained from neutron diffraction. ALTHOUGH some investigations of the crystal structures have been made of these four compounds previously,'-13 it was the purpose of the present investigation to expand the previous work in order to locate the various atoms, determine their coordinates, and to confirm or to correct some of the previous work. It was convenient to group these four compounds together since they are related chemicallv and/or structurally. The compound Ti2Cu is tetragonil; and Ti2Ni, Ti4Ni2O, and Ti4CU2O are all large fees of the same space group. Ti2Cu has been previously reported as a fee phase by Laves and Wallbaum;1 and Rostoker2 which was possibly the oxide phase, Ti4Cu20. Joukainen, Grant, and Floe;3 and Trzebiatowski, Berak, and Ramotow-ski4 have also reported a phase of this composition. karlsson5 has reported a small fct phase of the composition Ti3Cu which may be the presently discussed Ti2Cu phase. More recently Ence and Margolin6 have reported a small fct phase for Ti2Cu and the present authors7 together with Nevitt8 have briefly reported it to be a bet related to the fct with a co three times the length of the co of the fct and have also reported that this phase has a very limited composition. Further refinements will be reported which have varied some of the parameters of this bct structure slightly. Ti2Ni has been reported as a fee phase by Laves and wallbaum;1 Duwez and taylor;9 Rostoker;2 Poole and Hume-Rothery;10 and Yurko, Barton, and parr.11 In a later paper Yurko, Barton, and parr12 have given the complete structure of this phase based on an X-ray diffraction study which was independently confirmed with neutron diffraction by Mueller and knott.7 Additional crystal structure information will be given. Ti4Ti2O, Ti4Cu2O, and a number of other compounds including Ti4Fe2O have been reported as fcc phases by Rostoker,2 and more recently Nevitt13 has confirmed the Ti4Ti2O phase. Rostoker,2 however has reported diffraction lines for Ti4Fe2O which do not have all odd or all even indices. These lines, therefore, cannot be observed if this compound has a fee structure. This same error has crept into the diffraction results reported for TiNi2O and Ti4Cu20 in the ASTM powder data which has been credited from Rostoker's data. Complete crystal structures of these two phases will be presented. Although all four of these structures have large unit cells and hence do not lend themselves for completely resolved neutron powder patterns, a sufficient number of individual reflections was observed for solving the structure. They also serve as good examples of some of the advantages to be gained by using both neutron and X-ray diffraction techniques. EXPERIMENTAL PROCEDURE All of the alloys were prepared by arc melting. The starting metals had the following purity: Cu 99.999 pct, Ni 99.83 pct, and Ti 99.92 pct. Oxygen was introduced into the two oxide phases as chemically pure TiO2, with the remainder of the titanium coming from the above mentioned metal. All of the sample buttons were annealed in evacuated Vycor tubes, the two binary phases for 5 days at 700°C and the two oxide phases for 3 days at 900°C. Oxygen analyses were performed on all four phases by two independent laboratories with the following amounts of oxygen present in atomic percent; Ti2Cu-0.06, Ti2Ni-1.03, Ti4Ni2O-13.95, and Ti4Cu20-13.87. The stoichiometric amount for the oxide phases is 14.29 at. pct. Since all of the samples were very brittle they were easily reduced to a powder for diffraction measurements. The lattice constants given in Table I were determined for the four compounds from X-ray diffraction patterns of powder samples exposed to filtered copper radiation using a 114.59 mm diam Debye-Scherrer type camera using the Straumanis loading. None of the patterns showed a detectable amount of a second phase. The lattice constants were obtained from an IBM 704 computer program employing a least squares treatment with systematic correction terms as previously reported.14

Jan 1, 1963

THE following employment items are made available to AIME members on a nonprofit basis by the Engineering Societies Personnel Service, Inc , operating in cooperation with the Four Founder Societies Local offices of the Personnel Service are at 8 W 40th St , New York 18, 100 Farnsworth Ave, Detroit, 57 Post St , San Francisco, 84 E Randolph St , Chicago 1 Applicants should address all mail to the proper key numbers in care of the New York office and include 6c in stomps for forwarding and returning application The applicant agrees, if placed in a position by means of the Service, to pay the placement fee listed by the Service AIME members may secure a weekly bulletin of positions available for $3 50 a quarter, $12 a year MEN AVAILABLE Executive Engineer, 37, married, 3 children Now Resident Manager nonmetallic underground mine and processing plant Management and operational experience in gold, copper, zinc, limestone and lime products manufacture Mine plant design and construction, underground mines and quarries Excellent references Currently employed but available reasonable notice Prefer U S M-687-426-E-5-San Francisco Metallurgical Engineer, 29, married, one child, B S Metallurgical Engineering Five years' practical experience in milling and laboratory research Available thirty days Prefer West or Southwest U S A M-689¬522-E-1-San Francisco Experienced Mining Engineer, university graduate, 40, married Twenty years' experience all phases metal mining, past 13 years as manager, excellent cost and safety record Available short notice Best references and complete professional record available upon request M-690

Jan 1, 1952

By Reed W. Hyde

SOME information has been published on the sulphur dioxide concentration of gases from Dwight-Lloyd machines sintering lead ores but most of this relates to the customary practice in which the charge is relatively high in lead and is sintered in two passes over the machine. It is thought, therefore, that the following observations on sintering in one pass a low-lead, low-sulphur charge may be of interest. The data were obtained in sintering lead sulphide flotation concentrates mixed with minus 20-mesh oxide ore, the mixture carrying 16.8 per cent Pb and 5.3 per cent S. No returns were used because of the limited amount of sulphur available as fuel, but the mixture pelletized readily to make a satisfactorily permeable bed on the sintering machine. As there was a possibility that the SO2 might be utilized, a series of runs was made to determine the concentration of SO2 in the gases at different points along the wind box, and the effect on this concentration of changes in the wind-box vacuum. Several runs each were made with vacuums (as measured in inches of water at the wind box) of (1) 2 to 1 3/4 in., (2) 4 to 3 1/2 in. and (3) 6 to 5 1/8 in. The averages of the determinations from each series are platted on the graph (Fig. I). The vertical scale shows the percentage by volume of SO2 (as determined by the Reich method), and the horizontal scale shows the time expressed as percentage of the total sintering time. Sintering was considered to be complete when no SO2 could be detected in the gases. In all cases the depth of bed was 4 1/2 in. and the weight of dry charge per square foot of hearth area was 45.2 lb. Other data are given in the table. The volume of gases corresponding to the several points along the wind box at which SO2 determinations were made could not be measured, but has been calculated from the SO2 content and sulphur elimination on the assumption that the volume is uniform throughout the sintering period. This is not strictly true, as with a given vacuum the volume will usually drop slightly toward the middle of the period and increase toward the end. However, this assumption is a reasonable approximation. Several interesting features are brought out by these data. In general, the published data on SO2 concentration in sintering under ordinary conditions show a high SO2 concentration in the early stages followed by a rapid decrease in concentration and with relatively little SO2 in the gases from the last half or third of the operation. In other words, the greater part of the sulphur elimination is obtained in the first third or half of the period. In these runs, the 6-in. vacuum, which approximates the usual operating condition, gives a curve that follows this customary pattern. Maximum SO2 concentration is obtained 2 ½ min. after

Jan 1, 1942

By R. W. Armstrong, D. H. Feisel

Self-diffusion in pure crystalline solids has been described through extension of the Cohen and Tum-hull volume -fluctuation model originally proposed for diffusion in simple liquids. It is shown, for a number of crystalline solids, that the self-diffusion coefficient, D, obeys the relationship where a is the lattice spacing, v the Debye frequency, Z' the lattice-volume fluctuation per atom required for a diffusion jump, and (?V/Vo) the relative increase in the lattice volume resulting from thermal expansion between O°K and the temperature under consideration. The calculated value of Z' is dependent upon the difiusion mechanism and the lattice geometry. ThE need for an increased understanding of mass-transport processes in crystalline solids has resulted in much research to determine rates of self-diffusion.' For several metals (e.g., copper, silver, and iron), the experimental values of the self-diffusion coefficient, D, have been determined over a large interval of temperature and the accuracy of D has been investigated through repeated experiments. Also, a number of empirical and theoretical diffusion studies have established a rational basis for understanding the D values for the different elements.2"6 The empirical diffusion analysis by Sherby and simnad6 supplied the main stimulus for the present study. Sherby and Simnad, reviewing the self-diffusion data for a substantial number of pure crystalline solids, suggested that sufficient data were available to establish with some certainty the apparent intrinsic material properties which influence D. Without explicit regard for any mechanism of diffusion, these investigators obtained an empirical relationship for D which correlated the self- diffusion coefficient for each element with the melting temperature, crystal structure, and valence of the particular element. Their analysis is only unsatisfactory in that the effect on D of crystal structure is not very clear and the valence to be assigned to an element is not always straightforward. The present study is an attempt, from a very different viewpoint from that of Sherby and Simnad, to establish a basis for understanding the self-diffusion coefficients for crystalline solids. This analysis was undertaken for the purpose of developing a general expression for D which would explicitly depend on the mechanism of diffusion. The main result of the study is that diffusion in crystalline solids may be described in terms of a volume-fluctuation model which was originally proposed by Cohen and Turnbull7 to describe diffusion in simple liquids. The model appears to be useful for predicting self-diffusion coefficients in crystals. The following sections describe the volume-fluctuation model for diffusion, the application of this model to crystals, and the resultant agreement with experimental data. 1) THE VOLUME-FLUCTUATION MODEL FOR DIFFUSION Cohen and Turnbull7'8 developed their model for liquid self-diffusivity for the purpose of describing certain properties of liquids, glasses, and the liquid-glass transition. A hard-sphere model for the liquid structure is used. In this model, the size of a spherical molecule or atom is fixed for all temperatures by the interatomic spacing at the absolute zero of temperature. At any temperature above O°K, the atom is viewed as being contained within a larger spherical cage defined by the interatomic spacing of the atoms at that temperature. A random distribution of varying cage sizes is assumed to exist in the liquid. An individual diffusion jump occurs when the free volume of an atom (the volume of the cage containing the atom minus the volume of the atom) is large enough for the atom to jump out of its cage. With this physical picture of diffusion in liquids, Cohen and Turnbull derive the following equation for self-diffusivity: D=Doe-yv*/vf [1] where Do is the product of three factors which are the diffusion-path geometry, the diffusion-jump distance, and the velocity of the diffusing atom, v*

Jan 1, 1964

By J. B. Austin

IN considering so complex a process as the smelting of iron in the blast furnace, there is obviously no single method of calculating efficiency that gives a complete appraisal of the performance of the furnace in all its several functions; such a comprehensive view as is required to evaluate producing ability and manufacturing costs for a given furnace is obtained only by considering a group of efficiencies, each of which measures the performance of the furnace from a specific point of view. For instance, there is the rate of consumption of coke per ton of iron produced, which is probably the most frequently considered, and is certainly one of the most important of the group; there is the efficiency of recovery of iron charged; and there is the efficiency of utilization of energy, both chemical and thermal, in some respects related to the former two yet in other ways independent of either. In each of these cases, the calculation can be made on the basis of the actual input of material or energy, or it may be based . upon a comparison with the minimum amount of material or energy that would be required by a "perfect" furnace; that is, an imaginary ideal furnace analogous to the perfect steam engine used in thermodynamic calculations, when this furnace is operating under optimum conditions. Moreover, the furnace itself can be considered either as an apparatus for smelting iron or, taking a somewhat less common point of view, as a gas producer yielding iron and slag as by-products. Some of these effi-ciencies are more useful than others, yet all are instructive; for a knowl-edge of them not only leads to good engineering practice from the standpoint of control but also discloses limits to possible improvement in performance and indicates the lines along which improvement is most likely to be profitable. Estimates of some of these efficiencies have appeared from time to time in the literature, but many of them are open to objection because they are based on data of uncertain accuracy or because they have been calculated without due regard to the thermo-dynamic principles involved. Moreover, so far as the author is aware, no general survey of the blast furnace comparing its performance in all the aspects mentioned has been made.

Jan 1, 1938

By Colin J. Dixon

The full realization of the potential of computers in geological in- formation system demands new approaches to the machine representation of information. At the same time, the feasibility of such a system, particularly in commercial field, demands greater freedom in input preparation and a degree of standardization in the arrangement of subject matter system. The paper describes some of the basic notions of project "GEOSEMANTICA 70". The aim of the project is to devise means of storing real geological informations in a machine rather than the representation of the information in human language form. This is seen as the only long-term solution to the problems of the complex vocabulary of geology and the complex syntax of geological description. Effectively, a machine language is being defined, consisting of a set of symbols and a date structure for representation of geological and related information. Beyond this, is the definition of a set of rules allowing the use of constrained text as an input/output medium. The investigation is oriented primarily to the representation and handling of information about mineral deposits and the associated geographic, economic and engineering information.

Jan 1, 1969

By G. W. Phelps

The field of industrial minerals is concerned with both sieve size and distribution of sub-sieve particles. A great deal of work has been reported on the techniques designed to provide information of this work. The term sub-sieve is here considered to include particular material that will pass a 325-mesh testing sieve. Plant men require rapid, precise and simple testing techniques that are, if possible, inexpensive with respect to equipment and manpower. A number of particle-size procedures have been developed, based on the Stokes equation for streamline movement of a particle of diam D, moving at velocity V through a liquid of viscosity n. 1 Particles are considered to be rigid, smooth spheres, completely dispersed at low concentration and moving at low velocity through the suspending medium with no slippage between particle and liquid. In particle size determinations the particle moves either from natural gravitational forces or those provided by centrifugal means. The diagram of Fig. 1 shows a spherical particle of density S and diam D moving downward at velocity L/T (or V) through a liquid of density So and viscosity n. The forces of gravity G (natural) or Gc (centrifugal) cause the particle to move through the liquid, thereby creating a viscous drag F or Fc. Within the streamline flow

Jan 1, 1967

By Walter Bradley

MINING by hydraulic process of the important gold-bearing gravels of the Sacramento Valley in the basins of the Yuba, Bear and American rivers began in 1853, and continued at an ever-increasing rate for 30 years. In its most active days it reached an annual production of $15,000,000, and averaged $10,000,000 per year over the 30 years. During that period no provisions were made to prevent the debris from being dis-charged into the water courses, and ultimately large quantities of silt, sand and gravel were washed down to the lower valleys; filling channels, causing overflow of agricultural lands and difficulty in maintaining navigable depths in the Sacramento River. Agitation and suits filed because of these damages finally resulted in the injunction decision of Judge L. B. Sawyer,1 in the U.S. Circuit Court, Jan. 23, 1884, which closed down practically all hydraulic mining operations not having means to restrain their debris. In 1893 (approved March 1, 1893), Congress passed the act introduced by Representative Anthony Caminetti, of California, creating the California Debris Com-mission, composed of three officers assigned from the Engineer Corps of the U.S. Army, and since that time it has been legal and possible within the drainage systems of Sacramento and San Joaquin rivers to conduct hydraulic mining of gold-bearing gravels behind dams built under this commission's permit and supervision. In the watersheds of the Trinity and Klamath rivers, which flow directly to the Pacific Ocean and do not affect navigation or agriculture, operating permits from the California Debris Commission are not required. There is a measure of restriction there, however,. through what is known as the Trinity and Klamath River Fish and Game Dis-trict. Chap. 215, Stat. 1931, in order to secure relatively clear water for the fishing season, provides that between July 15 and October 15 each year there must not be discharged into these rivers debris that will cause a turbidity greater than a content of 50 parts per million in the

Jan 1, 1936

By Craig R. Barrett

The influence of the stress dependence of the dislocation density on the overall stress dependence of The steady-state creep rate is discussed. Experimental measurements of dislocation densities and creep rates jor an Fe-3.1 pct Si alloy tested over a wide range of stress are presented which indicate a close correlation between the two quantities. EXPRESSING the stress dependence of high-temperature steady-state creep rates has been the subject of numerous investigations over the past decade. From these investigations have emerged a few general equations which fit most of the available data reasonably well. Garofalo 1 has proposed that for a wide range of stress and temperature the creep rate, i, can be expressed by a relationship of the form: where A, ß, and n are temperature-dependent and is the applied stress. Eq. [I] gives a stress dependence of a at low stresses and exp(nßo) at high stresses, both consistent with experimental observation where n = 5. A number of other investigatorshave proposed a similar expression where: where m and w are in general functions of temperature. Theoretical justification for Eqs. 111 and [2J comes from two different approaches, theories based on dislocation glide3-5 and theories based on the balance between work-hardening and recovery.' Dislocation glide theories predict where p is the mobile dislocation density and V is the activation volume. In order for Eq. [3] to predict a on stress dependence at low stresses with n = 5, it is necessary for p to be a strong function of stress, about proportional to u3 or u4.436 Recovery models of creep start with the general expression:' where r is the recovery rate and h is the work-hardening rate, and then employ a detailed model to determine the stress dependencies of r and h. As example, weertman2 has described a model based on pile-up hardening which results in an expression similar to Eq. [2] with m = 2 and p = 2.5. A recent, seemingly more general approach suggested by McLean is to use Friedels dislocation network growth expression to calculate the stress dependence of r. This model predicts r = p 3,2 and, as h is taken to be a relatively weak function of stress, the stress dependence of i should be about proportional to p3, 2. Once again, it is required that p be a strong function of stress to yield i This recovery model is by no means complete as the stress dependence of network growth is calculated in the absence of an applied stress meaning that r should probably be a stronger function of stress than given by p3, 2 and also some recent work by Mitra and McLean8 suggests that the stress dependence of h may be as large as Because of the importance of the stress dependence of p in determining the overall stress dependence of the strain rate it seems reasonable to examine this relationship in some detail. Measurements made on dislocation densities present during high-temperature steady-state creep generally obey the following relationship a =o0 + aGVWp [5] where uo is usually called a friction stress, a is a numerical constant about equal to unity, G is the shear modulus, and b is the Burgers vector. Depending on the magnitude of uo, data obeying Eq. [5] can give rise to a range of stress dependencies of p. For example, if we assume that the expression where 6 is some constant, adequately represents the stress dependence of p over a range of o, then from Although Eq. [6] obviously breaks down when u 5uo for uo 6 can vary over a wide range of values. In particular, if u > a, > 0 then 6 will be greater than 2. As an example the data of Ishida and McLean are plotted in Fig. 1 showing excellent agreement with dynes per sq cm) and also showing good agreement with a p cc u3 dependence when plotted on a log-log basis. As oo increases 6 tends to increase and correspondingly the stress dependence of the predicted creep rate should also increase,* as-

Jan 1, 1968

By M. H. HARRINOTON

THE Transactions of the Institute afford abundant evidence of the general recognition by mining engineers of the importance of a safety-lamp which will not only give warning of the presence of fire-damp, but will also indicate the proportion in which it is present.' Mining-men are only too familiar with the frequent and sudden changes that constantly occur at the working-face of a gaseous mine. Nothing illustrates more forcibly the pregnant condition of the strata enfolding the seam and the seam itself, than the fact that a circulation of perhaps 150,000 cu. ft. of air may contain 2 or even 3 per cent. of marsh-gas, representing a volume of from 3,000 to 4,000 cu. ft. of this gas pouring-out each minute from the natural strata into the mine-workings. It would, indeed, be strange if this great outflow of gas were always constant, and it creates no surprise that such earth-breathings, as they may be called, are spasmodic and irregular in their wane and flow. There are other causes that act to vary the gaseous condition of mine-air. Gas has a tendency, often very manifest in mine-workings and airways, to stratify and, at times, to travel long distances in veins or streams. It is common to catch and lose the gas in the lamp when making a test. At times two lamps, held side by side in an air current, will indicate different gaseous conditions of the passing air. When testing for gas, the lamp may suddenly fill with flame, in a mine-chamber, where a moment before a careful test showed the presence of but 1 per cent. of gas. It is sometimes remarked that there is little need of detecting a smaller percentage of gas than is revealed by the common

Jan 1, 1906