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By Clyde E. Williams

MODERN steelmaking has gradually evolved from an inefficient small-scale operation, utilizing tiny units, to a highly efficient one utilizing large units almost completely mechanized. The leading position of the United States in the steel industry is due to the possession of easily assembled high-grade raw materials and large-scale highly developed plants. Modern integrated steel operation is largely one of material handling. Mining and transportation of iron ore, coal, and fluxes, coking the coal in by-product ovens, smelting the ore with coke and limestone in blast furnaces, converting the resulting pig iron into steel in the Bessemer, or in the open hearth when supplemented with scrap, and fabricating the resulting steel ingots in the shapes of commerce all are done on a huge scale, mostly by machinery, most of the labor being employed either in operating the machine or keeping it in repair. At the same time the quality and uniformity of the product have been greatly increased by the development of scientific knowledge and operating skills. Thus, the steel industry in this country is highly developed both mechanically and technically.

Jan 1, 1942

By Walter O. Roberts

A TOTAL eclipse of the sun is a brief, exciting spectacle witnessed by most men but once or twice during a lifetime. But to an astronomer an eclipse of the sun is an event of utmost scientific importance. For then, certain phenomena of the sun normally obscured by the brilliant glare of sunlight become spectacularly visible. For over seventy years prior to 1930 astronomers vainly attempted to discover a means of observing all of these eclipse features of the sun without a natural eclipse. But in 1930, through the genius of the French scientist, Bernard Lyot, astronomers met final success. Dr. Lyot invented a new kind of telescope known as a coronagraph, which made possible photography of all of the eclipse features of the sun. In 1940 Harvard University completed final testing of its Lyot-type coronagraph. the only one in the Western Hemisphere, and shipped it to Climax, where Donald H. Menzel, of Harvard College Observatory, supervised its establishment on the southeastern edge of the mining community of the Climax Molybdenum Co., with whose help and advice the modern residence and telescope dome of the observatory were constructed. Ever since that summer the station has been operated by me, first as a graduate student, and now as a staff member of Harvard College Observatory.

Jan 1, 1946

By Tracy Jarrett

In 14-karat gold alloys, as in 10-karat gold alloys,3 the addition of such metals as zinc, nickel, copper and silver produces a wide range of physical and chemical proper-ties such as color, hardness, corrosion resist-ance and melting points. These important factors made gold alloys desirable in the optical work and for jewelry. In 14-karat gold alloys, as well as in 10-karat or 12-karat, it is possible to have two golds of the same color, one of which will corrode more than the other, although it may be harder. Some of these alloys have been investigated by F. E. Carter,1 E. M. Wise,5 and L. Sterner Rainer.4 In this group of alloys the gold content of 58.3 per cent (14-karat) was held constant, while the copper varied from 36 to 24 per cent. The balance of the alloy consisted of one, two, or three of the metals silver, zinc, and nickel. All the alloys were melted in an induction furnace and cast into an ingot that meas-ured 3 by 1/4 by 3/4 in. The method of determining the melting points, color deter-minations and acid test have been discussed in previous papers by the author.2,3 It is necessary to use small test bars (Fig. r). The tensile bars were annealed at 1300°F. for 20 min. and then air-cooled. There being a very marked drop when the yield point was reached, "the drop of the beam" method was used in determining the yield point.

Jan 1, 1940

Ernest R.. WOAKES, Cana, Colombia: The allusion in this paper, under the heading " Mining on the Isthmus," to the Espiritu Santo mine at Cana is neither according to fact nor worthy of the most interesting and able summary of the mining districts of Colombia contained in the remainder of the paper. The following remark is certainly open to contention: "The most widely-known mine of this department is the Espiritu Santo, at Cana, a quartz-mine opened up in 1680 and abandoned in 1727 on account of the attacks of the Cana Indians. This mine undoubtedly produced great quantities of gold, although many of the accounts given of its ancient richness are probably exaggerated. In 1884 it was rediscovered and reopened by the Darien Gold Mining Company, and under the able management of Mr. Thomas H. Leggett, now prominent in South African mining circles, produced for a time successful results ; but later the veins became pockety and the output unsatisfactory." The mine here referred to, though situated at Cana, was not the Espiritu Santo mine, and was closed down in 1892. Mr. Leggett, however, ran ten stamps on the ore from it for some two or three months in the year 1890, when he left Colombia. The Espiritu Santo mine was not reopened by the Darien Gold Mining Company until the year 1894. For the last three years the ore from the old Spanish workings of this mine has been yielding in the mill 1 ounce to the ton, and has served to keep the mine going whilst extensive development-works have been

Jan 1, 1899

Economic Significance of High-grade Concentrates BY PAUL M TYLER AND CARLE R HAYWARD (Min & Met, Dec, 563 4000 words) Generally pyro¬mctallurgical methods yield a higher recovery but cost more per ton of ore treated Milling methods are inherently cheaper for the reason that they do not involve dissipa¬tion of heat energy in slag formation, ctc It is vaguely realized that the smelter man could do a better job if he were supplied with purer raw materials but hard¬headed mill men have been offered little inducement to do a really good job of cleaning concentrates To do this may involve some sacrifice in percentage yield and in mill capacity and unless there is a compensating saving in freight and smelter charges, it simply is not good business to go beyond a certain point in the way of making clean concentrate Opportunities exist for eliminating one or more steps in the expensive series of processes for recovering various metals, or at least reducing the volume of materials that has to be handled at the most costly steps The beneficiation of nonmetallic minerals likewise affords numerous examples of over-all economies obtainable through more refined methods of mechanical improvement All this leads to the conclusion that perfecting selective processes for isolating pure minerals (e g, galena) virtually free from other sulfides as well as gangue may be productive of industrial economies Conversely, cheapening smelting methods may invade the field of mechanical improvement of ores but, as Kiphng said, that is another story

Jan 1, 1937

By Carl Samans

IN their recent complete review of the subject of age-hardening,1 R. F. Mehl and L. K. Jetter classify the main types of precipitation-hardening alloys under two headings, depending upon the nature of the precipitate and the character of the precipitation. In the alloys falling under the first heading, which display "continuous precipitation" and are typified by the aluminum-copper alloys, "precipitation seems to occur without much regard to grain boundaries and apparently at much the same rate at every point," although an acceleration in regions plas-tically deformed has definitely been established .2 In the second type of alloys the precipitation is "discontinuous," as in the copper-silver system for example, and "seems to start in restricted regions, primarily those contiguous with the grain boundaries, and to spread inward in an advanc-ing wave." The reasons for the occurrence of these two greatly different types of precipitation are unknown, as are those for many other reactions that take place in the solid state. In the present paper only the first or "continuous" type of precipita-tion will be dealt with, for which a complete crystallographic mechanism has been worked out for the process of precipitation from a supersaturated solid solution of copper in aluminum. Based upon this mechanism, an explanation of the various property changes found experimentally will be developed in the second paper of this series. In later papers "dis-continuous precipitation" in copper-silver alloys will be explained, and an effort will be made to rationalize the two processes and types of precipitation.

Jan 1, 1940

In 1910, he was with the Hidalgo Guadalupe, Pachuca, Mexico, and in 1911 was engineer with the Gualcola Mines Co., at Tuquerres, Colombia. In 1913, he returned to the Sout.hwest, and was employed for that yea.r by the Inspiration Cons. Copper Co. at Miami, and in the middle of the following year he became engineer with the Detroit Copper Co., at Morenci. In the autumn of 1914, Mr. Reece, stirred by his English blood, resolved to return to England to take part in the war. At this time he wrote "I think it will be better for me to resign my membership. If I come through all right, I will apply for re-election." The Institute, however, instead of accepting Mr. Reece's resignation, voted him an indefinite leave of absence, and this precedent soon crystallized into the regular procedure, now in force, of remitting the dues of members in active service. On Dec. 17, 1914, he wrote from Montreal that he fully expected to go over shortly in the second contingent from Canada, with Borden's Armored Brigade, operating with armored motor cars. As he said "this will be active enough service to satisfy any enthusiast, as they are used generally for reconnaissance duty." Soren Ringlund Soren Ringlund died suddenly on July 24, 1918, at Fort Logan, Colo., where he was engaged in service with the Medical Department of the U. S. Army. Mr. Ringlund was born in Denmark in 1875. He came to the United States in 1902, and became a citizen in 1908. He graduated as

Jan 1, 1920

By D. N. Williams

THE present study was undertaken with two specific objectives: measurement of the thermal-expansion coefficient of ß titanium and examination of the effect of solid-solution alloying on the thermal expansion coefficient. Four titanium alloy niaterials were available in the form of annealed 5/8-in. diam rod. Three of these, Ti-8V, Ti-20V, and Ti-6A1-4V, were prepared in the laboratory from 140 Bhn sponge. The

Jan 1, 1962

By AIME AIME

AT the meeting on Engineering Education on Mon- A day afternoon E. A. Holbrook, of the University of Pittsburgh and chairman of the Committee, presided as chairman with W. B. Plank acting as vice- chairman. About sixty members of the educational fraternity were present. Dean Holbrook, after a few introductory remarks, introduced the first speaker of the afternoon, Charles Y. Clayton, who then presented his paper, "Why Do So Few Elect Metallurgy?" in which he stressed the fact that ,one of the things the study of metallurgy ,now needs is publicity and public education on what metallurgy is and 'does. Another thing he brought out is that the study of metallurgy has been looked on for many years merely as a branch of other departments. He gave tables showing past and present enrollment in the course of metallurgy in various schools.

Jan 1, 1930

By J. D. Verhoeven

The effectiue distriblltion coefficient has been Measured in a series of vertical, normal freezing experirtzents with large current densities passing through the solid-liquid interface. The results indicate that the effect of the current on solute redistribution is due to electvotransport in the liquid. The high currents produced extensive mixing in the liquid which was considerably reduced by solidifying in small lubing. Analysis of the convection indicates that it probably vesulls from a horizontal temperature gradient produced by the Joule heating- of the electvic crirver~t. Due to the strong dependence of this convection upon the specimen radius, it is conclnded that only for specit~ens with radii of less than somewhere between 1 and 2 mm wilt electric field freezing be getlerally effec-tire at enhancing the purification of zone-melting experittierats in loe-melting alloys. Experiments on 0.8-mm lubes illustrate some interesting potential applications of this techniqne as a means of solute control on smcrll specimens. In a number of papers published within the last few years'-4 it has been found that passage of large dc currents through the solid-liquid interface of a freezing alloy produces a change in the solute redistribution accompanying the solidification. It had been predicted theoretically by Hucke et al.s and by Pfann and Wagner that such an effect would occur as the result of electrotransport within the liquid boundary layer at the solid-liquid interface. These authors independently derived the following relation between the effective distribution coefficient and the variables of the process: Where k0 is the equilibrium distribution coefficient, V is the mobility of the solute relative to the solvent, E is the electric field intensity, D is the solute diffusion coefficient, R is the rate of interface motion, and 6 is the boundary layer thickness. Both groups of authors show that this equation predicts the possibility of extending the purification ability of zone-melting experiments by the addition of an electric current. However, there are other effects in addition to electro-transport which could be produced by the high dc current and might, perhaps, invalidate the applicability of Eq. [I]. For example, high electric current might affect the interface reaction or perhaps the equilibrium distribution which was assumed to exist at the interface. It was the purpose of this work to determine the applicability of Eq. [I] by systematically investigating a simple system for which all of the necessary parameters were known. It was of particular interest for speculating on the potential usefulness of this technique to determine the maximum practical field intensities obtainable, and the effect of these high fields upon convective mixing in the liquid. The initial work in this study was published in a previous paper3 which will be referred to as I. In that work, a current density of 2000 amps per sq cm was used with specimen diameters of 5 mm. The results gave a qualitative confirmation of Eq. [I.], and also showed that a large amount of convective mixing was produced in the liquid by the current. The present study is an extension of that work to smaller specimen diameters and higher current densities. Two points were overlooked in I and will be emphasized here: a) the value of ko has not been well-established in the Sn-Bi system and b) the observed convection could be due to thermal convection produced from the Joule heating. EVALUATION OF k0 To evaluate Eq. [ I.] it is necessary to know the value of the equilibrium distribution coefficient, k,. Tn I the value of k, was taken as 0.3 from a published phase diagram.7 According to ansen' the best data on the phase boundaries of the Sn-Bi system are those of Davidson.5 The data of Davidson give a k, value ranging from 0.21 to 0.25 for the compositions studied in this work, or an average value of around 0.23. In a recent publication, Rigaud and Tougas'10 determined ko by means of a solidification technique. For an alloy of approximately the same composition as that used in this work they obtained a ko value of 0.39. Hence, there is considerable uncertainty in the value of ko and, consequently, a series of experiments were made to determine k, more accurately.

Jan 1, 1968

By Ralph L. Erickson

The discovery of a Carlin-type gold deposit at Cortez, Nevada, in 1966 can be attributed directly to the use of geochemical exploration techniques. Most mineral deposits owe their discovery to geologic concepts, geologic analysis, and luck, but at Cortez, geochemistry as an exploration tool played the clearly dominant role in discovery. Of course, the economic significance of the discovery had to be determined by exploration and development drilling. The story of the discovery began in 1959 when field work was initiated on a new project, "Geochemical Halos Utah and Nevada," by R.L. Erickson and A. P. Marranzino for the US Geological Survey. The Cortez district was selected for geochemical work because it was an area with good geologic control where we could address the problem of how to prospect in barren outcrops for concealed ore deposits in potentially favorable structures (buried thrust zones) or favorable host rocks in the subsurface. Geologic mapping of the Cortez 15-min quadrangle, just being completed by Gilluly and Masursky (1965), showed that the quadrangle contained excellent exposures of both the upper and lower plates of the Roberts Mountains thrust fault, a major structural feature of north-central Nevada. Roberts (1960) had noted that a number of mining districts were associated with windows in the thrust. In 1959, Erickson and Marranzino did some reconnaissance rock sampling and spring-water sampling in the siliceous clastic rocks of the upper plate of the thrust. Results of this reconnaissance prompted a full-scale sampling program in 1960 in the upper plate rocks on the west flank of the Cortez window. Results of the investigation showed that anomalously high concentrations of metals occur in the upper plate rocks, and further, that the distribution of metals is fault controlled and shows a pronounced zoning pattern (central copper zone; intermediate zinc, copper, and lead zone; and outer arsenic zone). The anomalies were interpreted as primary leakage halos that originated from metal occurrences in the thrust zone or in carbonate rocks below the thrust and moved upward along normal faults that cut both upper and lower plate rocks. Erickson gave a talk about these anomalies in the summer of 1961 to the local AIME Section in Reno, Nevada; two short reports were published that year-" Geochemical Anomalies in the Upper Plate of the Roberts Thrust Near Cortez, Nevada" (Erickson et al., 1961) and "Hydrogeochemical Anomalies in Four Mile Canyon Near Cortez, Nevada" (Erickson and Marranzino, 1961). These releases prompted blanket staking in the area by several small companies. In 1963, geologic and geochemical mapping were started by the USGS in the lower plate carbonate rocks of the Cortez window west of the quartz monzonite stock at Mount Tenabo and north of the old townsite of Cortez. The results of the work showed anomalously high concentrations of arsenic, antimony, and tungsten in jasperoid, fracture filling, and shear zones in Silurian and Devonian carbonate rocks. The anomalous area was about 1.6 km (I mile) long and 300 m (1000 ft) wide. A brief report, "Geochemical Anomalies in the Lower Plate of the Roberts Thrust Near Cortez, Nevada" (Erickson et al., 1964a), was published in 1964. During this same time period, 1959-1964, several exploration or mining companies were active in the general area (chiefly mapping geology and acquiring property). In 1964, American Exploration and Mining Company (Amex) concluded that the entire Cortez district was worthy of an extensive exploration effort involving drilling as well as geologic, geophysical, and geochemical studies. To carry out the program, Amex formed a joint venture group with the Bunker Hill Company, Vernon Taylor, Jr., and Webb Resources. Their early efforts were directed to the upper plate rocks and to the lower levels of the old Cortez silver mine. The group also drilled some shallow rotary assessment holes adjacent to the area of the arsenic, antimony, tungsten anomaly in lower plate carbonate rocks described in Erickson et al. (1964a). Assay results from these holes offered little or no encouragement to the joint venture group. Splits of these drill samples were made available to the USGS. In order to enhance any metal content present in these rocks and to determine the mineral residence of any metals detected, heavy-mineral concentrates of each 3-m (10-ft) sam-

Jan 1, 1985

By Charles Legrand

In many iron, coal and copper mines where large tonnages are known before starting operation and proper provisions can be made, the problems of electric traction by trolley locomotives are not very different from those of surface plants. In such installations the gauge of the track, the radius of curves, and the clearances, both vertical and horizontal, can be made to suit the conditions of the traffic. It is more difficult to install electric traction in mines which were started with hand tramming and where no consideration was given to the possibility of mechanical traction being used. The writer having had some experience in the installation of electric traction in copper mines, the following remarks apply more particularly to these mines. With the gauge of track usually 18 or 20 in., the weight limit of locomotives obtainable from manufacturers in this country varies from 3 to 6 tons. The full-load speed varies from 49 to 6 miles per hour. These locomotives, being made to run on very small radius curves, have a short wheel base and a long overhang from axle to coupling, which necessitates a coupling' with a good deal of lateral motion to avoid derailing the cars on sharp curves, especially if couplings are of the standard railroad automatic type. Although 3-ton locomotives will run on 12- or 16-lb. rails, it has been found more satisfactory to use 25-lb. rails, as the track keeps in much better shape, it is easier to maintain the bonding in good order, and fewer derailments from dirt on the track occur with the larger rails. Where 6-ton locomotives are used the 25-lb. rails are satisfactory, but 40-lb. rails have proved cheaper where the traffic is heavy and the ground is soft, as the track maintenance is considerably lower with the heavy rails. The locomotives will run on 15-ft. radius curves, but on through runs it is advisable not to go below 40-ft. radius. The voltage used should not exceed 250 to 275 volts, and the trolley wire should be protected, to prevent accidental contacts, in front of chutes and at all points where it is low. With the air lacking somewhat in oxygen and the heat and high humidity prevalent in many mines, this

Jan 1, 1915

By L. S. Pratt

IN 1910 the author was engaged in a qualitative study of the radio-activity of several chemical substances and a few minerals. In the course of the work he studied the mineral allanite (obtained from granite ledges at Topsham, Maine), and found it to be appreciably radio-active. Recently the fact was brought to his attention that the radio-activity of allanite is not well known, although Hon. R. J. Strutt in 1905 published a list of radio-active minerals with their analyses, including allanite from Amherst County, Virginia. In the work done, a standard ,type of electroscope was used, and the readings were made with a telescope, fitted with a scale, by means of which the rate of discharge could be readily determined. The natural "leak" of the instrument was carefully determined, and all results corrected accordingly. The method of study was as follows: The material to be tested was finely powdered, and spread over the bottom of a small, inactive pan. The electroscope was then charged, and the pan placed in position. The door of the instrument was closed, and the rate. of discharge carefully determined. All results were reduced to common form to permit of comparison. It is to be noted that the materials tested were not used in strictly comparable quantities, such as if 1 g. of. each had been used, but about the same quantity of powder was used (in the cases of the solid substances), so that the results are qualitatively accurate.

Jan 5, 1916

By Harold C. Price

Prior to Sept. 1, 1928, there had never been constructed what might be termed a long pipe line with electric-welded field joints. A year later more than 2500 miles of electric-welded lines had been completed which varied in diameter from 6 to 20 in., and in length from 50 to 700 miles. This paper discusses methods used on five lines with a total length of about 750 miles. The welding equipment used for the firing line or rolling welds, was 300-amp. generators driven with six-cylinder gasoline engines, and for the bell-hole or tie-in welds, 200-amp. generators driven by four-cylinder gasoline engines. A welding gang consisted of a foreman, 6 to 12 firing-line welders and 2 to 4 bell-hole welders, a mechanic and a tractor driver. Coupons about 1/4 in. thick were frequently cut from completed welds, which showed a cross-section of the weld and an indication of its size and penetration. For tests of tensile strength complete welds were cut out. The tensile strength of 52 welds cut from 420 miles of 12-in. line averaged 50,870 Ib. per square inch. The following results give a fair average of the number of welds made per man per day: on 157 miles of 8-in. line both firing-line and hell-hole welds averaged 15.3—the firing-line average was 21; on 39 miles of 10-in. line the firing-line average was 19.9; on 50 miles of 6-in. line, with unfavorable weather, the average was 17.7; on 424 miles of 12-in. line most of which was welded during the spring, 10.3, and with good weather the average rose to 13 welds per day. The average time for making a complete weld on 12-in. pipe 0.313 in. thick was found to be 30 min. for the rolling weld and 45 min. for tie-in welds. A comparison of costs between electric and gas welding shows that the cost of labor for electric welding is 25 per cent. more than gas, generating cost 61 per cent. less than gas, welding wire 40 per cent. less than gas, and supervision and overhead the same. Estimating 50 per cent, annual depreciation on electric welding equipment, and no depreciation

Jan 1, 1930

By Joseph C. Mead, Alan Stanley

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus 5/8 in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9½ in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9½ in. bed sin-ters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite re-covery). These tests have proved successful in increasing ilmenite re-covery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was secured from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By H. E. Cline, E. F. Koch, J. L. Walter

It has been postulated and, in a few instances shown, that some kind of dislocation structure will be present at semicoherent interfaces to accommodate small lattice mismatches. In the present study of the NiAl-Cr eutectic, regular arrays of interface dislocations are observed at the boundary between the chromium-rich rods and the NiAl-rich matrix. The networks were examined by transmission electron microscopy and selected area diffraction. The rods and the matrix have a crystallographic relationship in which all directions and planes of the two phases are parallel. The dislocation networks are cmposed of a<100> dislocations lying on the intersections of the cylinders with (100) planes. Dislocations forming hexagonal rather than square arrays are also observed at certain areas of the network. The morphology of the network is consistent with the interpretation of mismatch being accommodated by interface dislocations in the cylindrical geometry. The measured spacing between dislocations was used to calculate an apparent lattice mismatch between the phases (˜0.35 PCt)interface network energy (-140 ergs per sq cm), and network strengthening (-10,000 psi). It has been proposed by Frank and Van der Merwe1 that dislocations should be present at the boundary between two semicoherent crystallographically related phases. The role of the interface dislocations would be to reduce the internal stresses, caused by the mismatch in atomic spacing across the interface. Such dislocations have been observed at the interface between expitaxially grown films and Substrates.2-4 Interfacial dislocations have also been observed at precipitate-matrix interfaces.&apos;-&apos; Directionally solidified eutectics have been shown to have semicoherent phases1&apos; and would, therefore, be expected to have interfacial dislocations as found by Weatherly at a lamellar fault in A1-A12Cu.11 The NiAl-Cr eutectic appeared to be a promising system to examine because both phases are cubic, the lattice mismatch is small, and the phases are crys-tallographically related. Furthermore, the eutectic is easily thinned for transmission microscopy. Indeed, interfacial dislocations were observed and this report describes the nature of the dislocation networks in the boundary between the NiAl-rich ma-trix-phase and the fine chromium-rich rods.21 I) EXPERIMENTAL PROCEDURE Ingots, 3/4 in. in diam and 6 in. long were made by melting in vacuum and casting under argon using 99.9 pct pure material. The composition, in at. pct, was 33 pct Ni, 33 pct Al, and 34 pct Cr. The ingots were then placed in A1203 crucibles on a water-cooled base, melted by means of induction, and withdrawn from the hot zone at the rate of 1 in per hr under argon.* * T his material was first directionally solidified in this laboratory by E. R. Stover Slices were taken perpendicular to the growth direction of the directionally solidified ingot for metallography and for transmission electron microscopy. The electron transmission samples were thinned mechanically, then thinned electrolytically in A-2 electrolyte* *A-2 electrolyte: 62 ml perchloric acid, 700 ml ethanol, 100 ml butylcellosolve, 137 ml distilled H20. until a hole appeared in the foil. 11) EXPERIMENTAL RESULTS A) Optical Microscopy. The microstructure, viewed on a plane perpendicular to the growth direction, is shown in Fig. 1. The structure consists of cells or colonies of parallel chromium-rich rods in the NiAl matrix. The cells occur when there are impurities present12 or, in a ternary eutectic, if the composition is slightly off the eutectic composition. The axis of the chromium-rich rods is parallel to the growth direction except near the cell boundaries. Here the rods may assume angles to the growth direction; however, examination shows that the crystallographic relationship between the rod and the matrix remains the same. Fig. 1 includes cell boundaries where the rods formed at a large angle to the growth direction. The variation of rod position across the cells made it possible to Fig. 1-Structure on plane perpendicular to growth direction. Rods near cell walls are at large angle to growth direction. Magnification 315 times.

Jan 1, 1970

By A. Stanley, J. C. Mead

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941,when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948. The magnetite concentrate produced in the milling operation was too fine (minus 20 mesh) to be used directly in iron blast furnace operation, and most of the magnetite had to be stockpiled in 1942 and 1943. In 1943, the Defense Plant Corp. built a Greenawalt sintering plant at Tahawus, N. Y., to put the magnetite concentrate in a more suitable form for use in the iron blast furnace. The Greenawalt sintering plant consists of three 10 by 25 ft sintering pans designed to produce 1800 gross tons of sinter per 24 hr. The vacuum to each pan is produced by two Greenawalt fans in series, pulling approximately 30,000 cu ft of air per minute at 50 in. water gauge vacuum. The plant started operation in August 1944. The present plant production averages 25 tons per operating pan hour (approximately 224 lb per operating hour per square foot of grate area) of plus 1 in. sinter. Raw feed to the plant consists of 61 pct magnetite, 4 pct anthracite coal culm, and 35 pct minus 1 in. return fines which are conveyed to a pug mill where the materials are mixed thoroughly and water added to give the mixture 5.5 to 6 pct moisture. The mixed prepared feed is conveyed to two 4 by 10 ft vibrating screens where the minus 1 in. plus 5/8 in. return fines are screened out and discharged into a surge bin for use as a hearth layer. The minus % in. prepared feed is discharged into another surge bin for use as prepared feed. A charge car, electrically operated, having a capacity of one charge of prepared feed and several charges of hearth layer, lays a thin layer of plus 5/8 in. return fines and 9 1/2 in. depth of prepared feed into the pans. A fluffing roll and a vibrator on the car fluffs and spreads the prepared feed into the pans. An ignition car, electrically operated, ignites the top of the bed with a 30 sec flash burn. The 9 1/2 in. bed sinters in approximately 13 min. Dumping the pan, and recharging and igniting the bed requires 2 min. To improve the quality of the ilmenite concentrate produced in the mill and to reduce the amount of titanium dioxide lost in the mill tailings and in the magnetite product, extensive research work and pilot plant operations have been done on grinding the crude ore to minus 65 mesh size (rather than to minus 20 mesh) and concentrating it by a combination of magnetic separation (for magnetite recovery) and flotation (for ilmenite recovery). These tests have proved successful in increasing ilmenite recovery and grade. With the development of the ilmenite flotation process to a stage where a full scale flotation plant was in the design stage, the problem arose of handling the 65 mesh magnetite concentrate that would be produced. In order to study and solve the problems of handling and sintering the 65 mesh magnetite in the sinter plant, a pilot sinter plant was plus from John E. Greenawalt. The effect of using 65 mesh magnetite in the sintering operations was then studied on the 2.4 sq ft test pan, operating under conditions as similar to the large plant as could be set up in the laboratory. A series of tests were run in the test pan on present sinter plant feed that had been mixed in the plant pug mill. An average production and an average quality of sinter produced in this series

Jan 1, 1950

By R. W. Raymond

R. W. Raymond, New York City: In his paper at the New York meeting (Trans., xix., 28), Mr. Ledyard described, among other mines, the Belmont, in the county of Peterboro, Ontario, the developments of which at that time indicated, though they did not, perhaps, completely demonstrate, a deposit of high-grade Bessemer magnetite of very great extent. Mr. Ledpard said that experts estimated the quantity of ore within 100 feet of the surface at 1,000,000 tons. But the precise data upon which this estimate was based were not furnished. We are all interested in the development of fresh supplies of Bessemer ores. I think nobody who knows how enormous is the demand can seriously fear that such deposits will be exploited too rapidly. On the other hand, I need not here enlarge on the fact that a considerable number of enterprises, in which American ironmasters have been more or less interested, for the development of Canadian iron-ore deposits, have been unsuccessful, more by reason of disappointment as to the quantity of the ore and the facilities for cheaply mining it on a large scale, than as to its quality. Of course, a Canadian nine supplying the American market is handicapped at present, to the extent of 75 cents per ton, by the duty on foreign iron-ore; but this would certainly not be prohibitory in the case of a deposit sufficiently large, rich, and accessible. Or, to put it otherwise, there might easily be, somewhere in Canada, near our borders, a mine so favored that its product could be profitably

Jan 1, 1892

[NUMERICAL CODE TO MAPS (Region Listed in Roman Numerals after Each Petroleum Section) 1 Alaska 2 Arizona 3 Black Hills 4 Boston 5 Carlsbad Potash 6 Central Appalachia 7 Chicago 8 Cleveland 9 Colorado 10 Columbia 11 Connecticut 12 Delta (VIII) 13 Detroit 14 East Texas (VI) 15 El Paso 16 Gulf Coast (VII) 17 Kansas (IX) 18 Lehigh Volley 19 Mid-Continent (IX) 20 Minnesota 21 Montana 22 Nevada 23 New York]

Jan 1, 1961

By J. F. K. Brown, E. E. Roecker

For three years The Hudson Coal Co. has used egg anthracite instead of coke in its foundry cupola. It has long passed the stage of being told it cannot be done—the metal would be cold, of poor quality, or the speed of operation would be low. From observation, from use and from test. we know that our metal is about average; the cast iron invokes no adverse comment from the machine shop and the speed suits our foundry requirements. We have there-iore resolved into practice anthracite combustion processes in an industrial furnace formerly using coke, under conditions suitable to our needs. When it comes to translating what we know into use in another cupola, however, where operating conditions may be different, or to explaining why, with seemingly the same conditions in our own cupola, we have attained sometimes a temperature of 2800° in the metal and at another time 2600°, or a speed of 5 tons per hour against 2 35 tons per hour, we cannot do it. We cannot do this because there seems to be an unbridged gap between laboratory ideas on combustion and practical ideas on com-bustiqn, which we have been unable to cross in terms understandable to us as practical consumers. This gap or bridge centers around the problem of how combustion actually takes place; and we believe it is controlled by two factors: (I) the character of the surface of the fuel and (2) the space provided and needed within which the oxygen of the air is allowed to or can act on that surface. It has long been contended on behalf of coke that porosity is an important factor in combustion, although, admittedly, there are two schools of opposing thought on this question, as revealed in technical literature. It has been understood generally that a light, porous coke is the best metallurgical fuel because there are more square inches of carbon exposed to the blast of air, therefore more useful carbon-dioxide reactions can take place in the given space between the entrance of the air and the point where useful work is done. It is now our general idea that the heavier and the harder the coal, the better the metallurgical fuel, provided the surface of that fuel is suitable for fast reaction with the oxygen. Time and Space A cupola is loaded in layers of fuel and iron. The horizontal line where the first charge of iron rests on the coal is the top of the bed coal. With us that is 22 in., usually, above the tuyeres where the air enters. This means 600 lb. of coal. Every subsequent coal layer is 100 to 125 lb. of coal, and that 100 to 125 lb. of coal melts 1000 lb. of iron at around 2700° in the ladle. This quantity of coal represents a 3-in. layer spread across the cupola. The question arises: If this 3 in. of coal will melt 1000 lb. of metal, why was it necessary to use 600 lb., or 22 in., of coal beneath and between the first charge of iron and the air entrance?

Jan 1, 1944