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William B. Cogswell, member of the Institute since 1872, died on June 7, 1921, at his home in New York City, after an illness of about six weeks occasioned by an infection of the middle car. Mr. Cogswell is survived by his wife, Mrs. Cora Brownell Cogswell; a son-in-law, and three grandchildren. His first wife died in 1877. William Cogswell was born at Oswego, Sept. 22, 1834, and was taken to Syracuse by his parents when he was four years old. He attended Hamilton Academy and private schools in Syracuse and Seneca Falls. In 1843, he visited England with his parents and brought back a set of drawing instruments, which led to a series of lessons in architecture under Luther Gifford, the only architect in Syracuse at the time. The boy drew many building plans and one set was used for the old Globe Hotel in Syracuse, erected in 1846 and 1847. In 1848, when only 14, he spent a year with a party that was surveying the route of the Syracuse & Oswego R. R. and relaying the tracks of the Utica & Syracuse R. R. He studied civil engineering at the Rensselaer Polytechnic Institute, Troy, for three years, but left that institution in 1852 without receiving his degree. Many years later, in 1884, the degree of C. E. was conferred upon him. During the three years following 1852, Mr. Cogswell served as an apprentice in machine shops at Lawrence, Mass., under the superintendence of John S. Hoadley. About 1856, he was made manager of the machinery department of the Marietta & Cincinnati R. R., at Chili-cothe, Ohio, and three years later became superintendent of the Broadway foundry at St. Louis, Mo. In 1860, in Syracuse, Mr. Cogswell was associated with William A. and A. Avery Sweet in the formation of Sweet Bros. & Co., which later became the Whitman & Barnes Mfg. Co. At the outbreak of the Civil War, Mr. Cogswell received a civilian appointment as mechanical engineer in the United States Navy. During 1860 he was located at Port Royal, S. C., having general supervision over the work of fitting up repair shops at five widely separated stations on the Atlantic seaboard and Gulf of Mexico. In this year his genius and resourcefulness were notably shown. He literally launched a naval machine shop, so that warships could be repaired without leaving, and thereby weakening, blockades against Southern ports. The machine shop was assembled by Mr. Cogswell and shipped by boat to Port Royal. There an old whaler was made over for machine shop purposes and Mr.

Jan 1, 1922

William B. Cogswell, member of the Institute since 1872, died on June 7, 1921, at his home in New York City, after an illness of about six weeks occasioned by an infection of the middle car. Mr. Cogswell is survived by his wife, Mrs. Cora Brownell Cogswell; a son-in-law, and three grandchildren. His first wife died in 1877. William Cogswell was born at Oswego, Sept. 22, 1834, and was taken to Syracuse by his parents when he was four years old. He attended Hamilton Academy and private schools in Syracuse and Seneca Falls. In 1843, he visited England with his parents and brought back a set of drawing instruments, which led to a series of lessons in architecture under Luther Gifford, the only architect in Syracuse at the time. The boy drew many building plans and one set was used for the old Globe Hotel in Syracuse, erected in 1846 and 1847. In 1848, when only 14, he spent a year with a party that was surveying the route of the Syracuse & Oswego R. R. and relaying the tracks of the Utica & Syracuse R. R. He studied civil engineering at the Rensselaer Polytechnic Institute, Troy, for three years, but left that institution in 1852 without receiving his degree. Many years later, in 1884, the degree of C. E. was conferred upon him. During the three years following 1852, Mr. Cogswell served as an apprentice in machine shops at Lawrence, Mass., under the superintendence of John S. Hoadley. About 1856, he was made manager of the machinery department of the Marietta & Cincinnati R. R., at Chili-cothe, Ohio, and three years later became superintendent of the Broadway foundry at St. Louis, Mo. In 1860, in Syracuse, Mr. Cogswell was associated with William A. and A. Avery Sweet in the formation of Sweet Bros. & Co., which later became the Whitman & Barnes Mfg. Co. At the outbreak of the Civil War, Mr. Cogswell received a civilian appointment as mechanical engineer in the United States Navy. During 1860 he was located at Port Royal, S. C., having general supervision over the work of fitting up repair shops at five widely separated stations on the Atlantic seaboard and Gulf of Mexico. In this year his genius and resourcefulness were notably shown. He literally launched a naval machine shop, so that warships could be repaired without leaving, and thereby weakening, blockades against Southern ports. The machine shop was assembled by Mr. Cogswell and shipped by boat to Port Royal. There an old whaler was made over for machine shop purposes and Mr.

Jan 1, 1922

The Pinson Mining Co., which will start production in very early 1981, is the fourth Carbon-In-Pulp plant to be put into operation in the United States. The property is located in Humbolt County, Nevada, 38 km (24 miles) northeast of the town of Golconda. The famous Getchell Mine was also near Golconda and was the site of one of the early Carbon-In-Pulp plants in the 1940s. The deposit was discovered in 1971 by Reno based Cordilleran Explorations, which is founded by a consortium of Canadian mining companies. Cordilleran is a partnership consisting of John Livermore and Peter Galli. In 1972 a second but smaller deposit, called "Preble", was discovered at a location 15 miles south of Pinson. During 1980, a decision was made to place the Pinson deposit into production and Pinson Mining Company was formed. Ownership of Pinson is vested in Lacana Mining Inc. , Rayrock Mines, Inc. , Siscoe Metals, Inc., John S. Livermore, Peter E. Galli and Donald M. Duncan. Three distinct but connected ore zones occur at Pinson. Combined reserves of milling ore total 3, 200, 000 tons at an average (undiluted) grade of 4. 12 g per mt (0. 12 oz per st) Au. Initial mining will be from "A" zone which contains 1, 500, 000 tons with an average (undiluted) grade of 5. 8 g per mt (0. 17 oz per st) Au. An additional 5, 500, 000 tons of low grade ore will be mined from the Pinson deposit and processed by heap leaching. Mining reserves at Preble total at least 1, 500, 000 tons grading 2. 7 g per mt (0. 08 oz per st) Au. Production plans for the Preble deposit have not been formulated at the date of writing. Gold mineralization at both deposits is finely disseminated as sub micron sized particles. The ore at Pinson is structurally controlled and occurs in an Ordovician limestone-siltstone unit. Much of the ore has been silicified to a jasperoid. The ore zones have been entirely oxidized. Preble ore occurs in the Cambrian formation in a carbonaceous, limey shale. Only the upper part of the Preble is oxidized and amenable to conventional cyanide leaching. The Pinson ore body is adjacent to the mouth of Granite Creek in the Osgood Mountains at an elevation of 1, 615 m (5, 300 ft) with the mill being 1. 2 km (3/4 mile) from the mine at an elevation of 1, 495 m (4, 900 ft). Mine development work began in May of 1980 with ore being stockpiled for later processing. Pertinent factors affecting

Jan 1, 1981

By Frank H. Macpherson

When it became apparent early in 1941 that the United States might be drawn into the war, studies were made of the bauxite situation in Arkansas, principally because 9.5 pct of the known bauxite reserves of the country were located in that State. Bauxite mining was then geared for maximum production to supply ore to meet this country's war requirements and, if necessary, to supply the needs of Canada as well. In view of the comparatively small bauxite reserves, Congress appropriated funds for an extensive exploratory drilling program to be carried out by the Bureau of Mines in cooperation with the Geological Survey. An alumina plant was erected near Bauxite and an aluminum plant near Hot Springs, Ark., by the Defence Plant Corp. The Metals Reserve Co. established a schedule of prices, arranged for stockpile sites and a buying agency. Operators cooperated to step up production and new ones came into the field. Production Record Prior to 1941, the greatest annual production in Arkansas was 562,892 long tons, dry basis in 1918. During the single month of August 1943, 693,596 long tons were mined—130,704 tons more than for the entire year of 1918. In general the Arkansas bauxite deposits are relatively small, irregular in thickness, somewhat scattered and, compared to coal, iron or open-pit copper deposits, are ill-adapted to an expansion in output. Commercial deposits range in horizontal extent from less than an acre to about 40 acres and are usually very irregular in outline. The average thickness of all known workable deposits is between 12 and 15 ft. with a maximum of about 60 ft. The ore bodies are usually somewhat undulating and nearly horizontal although locally some deposits dip as much as IS°. Types of Bauxite Three physical types of bauxite may be recognized: pisolitic, granitic and amorphous. Pisolitic ore is the most common type. It consists of small pebble-like nodules, ranging from the size of a pin head up to several inches in diameter, all set in a fine grained matrix. Transported bauxite is usually- pisolitic. Most residual bauxite is granitic. It is believed not to have moved appreciably from the location where it originally was formed. This is indicated by the fact that it appears to grade into unaltered nephelite syenite, which it overlies. Amorphous ore is the least abundant. It is nonpisolitic and resembles clay, has no characteristic texture, and can be identified definitely only by chemical analysis. Uses of Bauxite Bauxite is chiefly used for, and at the present time is substantially the only

Jan 1, 1948

By M. Smutz, C. J. Baroch, E. H. Olson

IN 1949 an orebody containing some 10 billion lb of recoverable rare earth metals was discovered in the Mountain Pass district of San Bernardino County, California.' The following year Molybdenum Corp. of America purchased a number of claims at the deposit and a mill on the property which had been constructed to process the district's gold ores. Average composition of ore from the San Bernardino deposit is less than 0.1 pct Tho2, 25 to 35 pct calcite, 10 pct rare earth oxides, 15 to 20 pct silica, and 30 to 40 pct barite.' Rare earth content of the ore analyzes 50.7 pct CeO2, 4.2 pct Pr6O11, 11.7 pct Nd2O3;, 1.3 pct Sm2O3,, and 34.3 pct La2O3 and others.' By floating the barite and depressing the bastna-site, Molybdenum Corp. of America produced a concentrate that contained 60 to 70 pct rare earth oxides. More than 25 pct of the rare earths, however, was lost in flotation. It seemed likely that a more economical method could be found. At the present time most of the rare earths in the concentrates are converted to rare earth chlorides, which are either fed to ion exchange columns for separation of the individual rare earths or reduced to misch metal. Current investigation shows that solvent extraction may achieve this separation more economically than ion exchange. The study reported here was undertaken to develop an economical method of preparing a pure rare earth nitrate mixture from California bastnasite ore. Results indicated that more than 98 pct of the rare earths could be recovered. Several possibilities were considered for processing bastnasite ore, but only the method outlined in Fig. 1 was investigated. This involved leaching the calcined ore with nitric acid. The nitric acid process is discussed under the following headings: 1) calcination of the ore, 2) leaching of calcine, 3) filtration studies, 4) solvent extraction, and 5) nitric acid recovery. Calcination: Since California bastnasite ore contains 25 to 35 pct calcite and about 10 pct bastnasite, calcination was needed to reduce the violence of reaction with nitric acid. Weight losses of 26 to 30 pct were obtained by calcining the ore at 900°C for I hr or at 800°C for 4 hr. Longer periods resulted in no further weight loss. Leaching the Calcine: When the bastnasite calcine was leached with 30 pct nitric acid for 8 to 10 days, rare earth recoveries of only 45 to 50 pct were

Jan 1, 1960

JAMES E. HOWARD,* Washington, D. C. (written discussion?). -Mr. Comstock, in his classification of the possible causes of transverse fissures, has not quite caught the point of view of the writer. It appears, following the trend of thought in Mr. Comstock's paper, that three groups instead of two might bee enumerated, namely: Fatigue of the steel, quality or, preferably, grade of steel, and mill practice. All steels are liable to failure by fatigue. Stresses that are capable of causing fatigue fractures are modified in their magnitude by the grade of steel acted upon. Mill practice is a term having a very wide meaning and in its relations to the development of transverse fissures has been used with such vagueness as to include many known factors and some that are unknown in the arts. According to the grade of steel employed, fatigue fractures may be accomplished by loads of greater or less magnitude, hence there is a relation between quality, or grade, of steel and the display of fatigue fractures. Mill practice of such an order as to affect the formation of transverse fissures must be shown to exert an influence on the ability of the steel to endure repeated stresses. Different grades of steel are selected for different purposes. Steel employed for fire-box purposes is of different grade from that used for cutting tools; although each grade is suitable for its purpose, and therefore of good quality. It is conceivable that a demand might be made for steel of physical properties in excess of those of any known grade. Under such a demand, impossible of' fulfillment, it would not be proper to refer to the steels as being deficient in quality; to do so would be a failure to recognize that all steels have their limitations in strength. Service conditions peculiar to rails admit of successive advance until the conditions can no longer-be met. The failure of a large number of rails from one type of failure, when all known means of averting the failure have been employed without success, is sufficient evidence that the end is close at hand. Iron rails, for a time, successfully met the requirements of railway service but weight of equipment, motive power, and rolling stock were increased, necessitating the use of steel. The early steel rails enjoyed the reputation of being of excellent quality, although: neither chemically nor structurally would they be so regarded ,lit the present time. They did good service in their day and under the conditions then prevailing were good rails. To meet the increased weight of rolling stock there has been

Jan 3, 1919

By Frank H. Macpherson

When it became apparent early in 1941 that the United States might be drawn into the war, studies were made of the bauxite situation in Arkansas, principally because 9.5 pct of the known bauxite reserves of the country were located in that State. Bauxite mining was then geared for maximum production to supply ore to meet this country's war requirements and, if necessary, to supply the needs of Canada as well. In view of the comparatively small bauxite reserves, Congress appropriated funds for an extensive exploratory drilling program to be carried out by the Bureau of Mines in cooperation with the Geological Survey. An alumina plant was erected near Bauxite and an aluminum plant near Hot Springs, Ark., by the Defence Plant Corp. The Metals Reserve Co. established a schedule of prices, arranged for stockpile sites and a buying agency. Operators cooperated to step up production and new ones came into the field. Production Record Prior to 1941, the greatest annual production in Arkansas was 562,892 long tons, dry basis in 1918. During the single month of August 1943, 693,596 long tons were mined—130,704 tons more than for the entire year of 1918. In general the Arkansas bauxite deposits are relatively small, irregular in thickness, somewhat scattered and, compared to coal, iron or open-pit copper deposits, are ill-adapted to an expansion in output. Commercial deposits range in horizontal extent from less than an acre to about 40 acres and are usually very irregular in outline. The average thickness of all known workable deposits is between 12 and 15 ft. with a maximum of about 60 ft. The ore bodies are usually somewhat undulating and nearly horizontal although locally some deposits dip as much as IS°. Types of Bauxite Three physical types of bauxite may be recognized: pisolitic, granitic and amorphous. Pisolitic ore is the most common type. It consists of small pebble-like nodules, ranging from the size of a pin head up to several inches in diameter, all set in a fine grained matrix. Transported bauxite is usually- pisolitic. Most residual bauxite is granitic. It is believed not to have moved appreciably from the location where it originally was formed. This is indicated by the fact that it appears to grade into unaltered nephelite syenite, which it overlies. Amorphous ore is the least abundant. It is nonpisolitic and resembles clay, has no characteristic texture, and can be identified definitely only by chemical analysis. Uses of Bauxite Bauxite is chiefly used for, and at the present time is substantially the only

Jan 1, 1948

Located 19 km (12 miles) north of the Carlin gold mine and 43 km (27 miles) north of Carlin, Nevada, is Carlin Gold Mining Company's Bootstrap mine. This deposit is similar in character to the Carlin ore, except it occurs in the upper plate formation of the Roberts thrust fault rather than in the lower plate rocks. The deposit was discovered over 50 years ago and operated on a small scale in the early 1960s using vat cyanide leaching. Newmont Mining Corporation acquired the property in 1967, and from an open pit mined 650,000 mt (717,000 st) of ore which was trucked to the Carlin conventional cyanide plant. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME, 1975. ) With the same grade of ore, the gold recovery by conventional methods is somewhat lower than that of the Carlin ore, even with a finer grind. Leaching tests were first conducted by the United States Bureau of Mines in 1968; and in 1971, ‘72, Carlin conducted extensive cyanidation investigations by bottle testing, simulated heap leaching in tall columns, and small scale pad heap leaches. Gold extractions ranged from 55% to 95% with low reagent consumptions. In 1974, a 545 mt (600 st) heap leach test was made at Carlin on pit-run low-grade material assaying 1.37 g per mt (0.04 oz per st). In a six week period the extraction of gold was 47.57'0. Carlin uses atomic absorption analyses on cyanide leach liquors to determine potential leach ore grade rather than fire assaying because it gives a more practical assessment of leach recovery than does fire assaying, which accounts for the total gold content. Before open-pit mining was started, plans were made to heap leach, the ore at Bootstrap which was too low grade for transport to Carlin. The topsoil outside of the designed pit area overlaid ancient lakebed clays which were found suitable for compaction to form an impervious base on which the dumps to be leached could be placed. By using the cost of truck haulage and milling at the Carlin plant and the price of gold 'at that time, a cutoff grade was established which resulted in the gold content of the first 150,000 mt (165,000 st) of the heap leaching ore being 1.51 g per mt (0.044 oz per st). An additional 500,000 mt (552,000 st) of lower grade material, assaying 0.96 g per mt (0.028 oz per st), is available for leaching as economics permit.

Jan 1, 1981

By B. M. Larsen, T. E. Brower

THIS discussion is based on statistical manipulation and evaluation of operating data from several commercial blast furnaces which include rather wide variations in practice. We are concerned here mainly with the effect of manganese on sulphur elimination under operating conditions in commercial blast furnaces, but this effect can be outlined more clearly by including certain other variables such as silicon in iron and slag basicity, which are all in part interrelated. Real progress in understanding all of the smelting-zone reactions in this process requires more laboratory studies of slag-metal melts similar to those in the blast furnace, with careful control of oxygen pressure. Such experimental work is very difficult, however, and in the meantime, with the present scarcity of more fundamental data, such statistical trends and correlations as can be obtained from commercial furnace operating records may help in choosing laboratory test conditions as well as adding to our knowledge of operating procedures. We do not know just why the furnace data are apparently so irregular that statistical treatment of large numbers is needed to obtain accurate trends. It is partly caused, no doubt, by the many simultaneous variables present, but there are perhaps also such factors as: (I) the con- tinual swings up and down so characteristic of the process; (2) difficulty of accurate sampling of co-existing phases, as well as (3) the presence of both equilibrium and rate factors as a result of the fast driving of the process. In desulphurization, for example, the flow of thin layers of slag and metal over solid coke surfaces, plus the contact between droplets of iron falling through the slag layer in the hearth, would lead one to expect a close approach to equilibrium. On the other hand, most of the critical part of sulphur removal may occur in a short time and distance below the tuyeres with perhaps an accompanying decrease in oxygen pressure and the concurrent reduction of much of the silicon and manganese, so its degree of approach to equilibrium may be quite variable. Probable Chemistry of Desulphurization Since there is little evidence that sulphur atoms in the iron solution are aasociated to any appreciable extent in molecules such as MnS, the transfer from slag to metal and back again can be written:* (FeS) ? [FeS], [I] hence the following reactions seem most probable in the slag phase, (FeS) + (CaO) ? (CaS) + (FeO), [2] and, (FeS) + (MnO) ? (MnS) + (FeO). 131

Jan 1, 1949

By Roy C. Kirkman

Dear Editor: It is with great pleasure that I read Mr. Howard M. Wells article in the December 1978 issue of MINING ENGINEERING entitled: "Optimization of Mining Engineering Design in Mineral Valuation." Mr. Wells very adequately presents the case and point that there are a great number of variables which must be analyzed during an economic evaluation of an FUND RESOURCE, such as copper or uranium. I wish to take this opportunity to suggest that, given today's economic climate, decision making takes place under conditions of uncertainty and that this uncertainty justifies our need to examine each one of the parameters entered into the evaluation model. It is my contention that the influence of the tonnages and grades as well as their distributions, have the most profound effect upon the results of an evaluation. Everyone concerned with the evaluation of FUND RESOURCES realizes that as metal prices decline or stagnate with sharply rising direct operating costs, cutoff grades (Gc) must increase according to the equation below which Mr. Wells uses: Gc = Total Operating Cost/(Market Price X 20 X Recovery) However, historically, we have seen cutoffs move dramatically lower in both the copper and uranium industries for the past decade. Henceforth, this established the metals industry most important argument for processing greater tonnages of lower grade materials from disseminated deposits (refer to paper by Kirkman, Roy C.). These disseminated deposits, created by either concentration processes (uranium) or dispersion processes (copper, gold, molybdenum), have log-normal grade distributions (Fig 1). Gross profits are rising, but unit profits are falling. The industry has created longevity through economy of scale, i.e., mining vast tonnages of low grade. In order to develop the great tonnages necessary to prove economic viability, we must examine alternatives at the low-end of the tonnage grade relationship. Notice from [Fig. 2] that as we move further to the right away from line A-B drawn tangential to the average grade curve at 45°, total profits will increase (given fixed annual production capacity) and unit profits decrease. The theoretical optimum operating point is where: [ ] At this point my observation is that slight shifts in production capacity bring about greater changes in profitability due to greater grade changes. Therefore, the best financial risk evaluation is obtained by totally revising the mining engineers methods and fixing the financial parameters such as discount and interest rates. A 10% shift in either of these two parameters will not cause near the same variance in economic viability as will a change on the GRADE-TONNAGE curve [(Fig. 2)].

Jan 1, 1980

By Edward L. Fidler, Thomas E. Finch

Thin seam coal mining in the western US nominally equates to recovery of stray seams associated with thicker, major seams. The thin or stray seams encountered are generally recovered, but not in all cases. The alternative to extraction is, of course, spoiling. The majority of thin seams are encountered somewhere in the overburden above the major seam or in the parting between thick seams. Although some are true strays and appear anywhere in the geologic sequence, others might be classified as thin "rider" seams immediately above or below a thick seam and separated by some equally thin parting. For the purpose of this discussion, stray or thin seams are defined as those which are less than 1.2 m thick or represent less than 10% of the expected extractable coal. [Figure 1] is a graphic representation of typical thin, stray, or rider seams which actually appear in a western operation. Operationally, stray seams create irritating scheduling problems and require special handling techniques which do not always lead to efficient equipment utilization. Additionally, coal recovery from a thin seam is considerably less than the 95% usually recovered from thick seams. Higher recovery in thin seams leads to dilution problems, which depreciates Btu quality, and most western coals are not in a position to lose heating value. Thin western seams also undulate and pinch or swell more than thick seams, and this leads to more difficult recovery. Often they are associated with equally dark carbonaceous shales and visual selectivity is difficult at best during daylight hours. Resource Recovery Regulations Thin seam recovery in western mines is encouraged by both federal and state regulations and by the mining industry's concern for total resource recovery. [Table 1] lists the western coal states and their recovery regulations. The federal regulations are found in the Surface Mining Control and Reclamation Act of 1977. They state, "General performance standards shall . . . require the operation as a minimum to … conduct the surface coal mining operations so as to maximize the utilization and conservation of the solid fuel resource being recovered." Economic stimulation for thin seam recovery appears to be marginal or at the least controversial. Scheduling and operational problems often outweigh the nebulous profits created by engineering analysis. The only alternative to extracting thin seams occurring above major production seams is to strip directly through them and lose the resource to the spoil pile. Although this is not palatable to environmental emotions or resource promoters, it is the most simple expedient to operations people.

Jan 1, 1981

By P. Farrar, H. Margolin

The Ti-Pb diagram was investigated in the region 0 to 58 pct Pb and from 500°C to liquidus temperatures. Three reactions were encountered: I—ß?a+Ti Pb at 725 10°C; 2—ß+L?Ti4Pb at 1305+ 10°C; and 3—the possible eutectic L-+Ti,Pb and y between 1200" and 1300°C. LEAD-RICH alloys have been investigated by Nowotny and Pesl,' who used sintered compacts and studied the region from 37.5 to 100 pct Pb by X-ray diffraction analysis. The only phase found in addition to titanium and lead was the hexagonal compound Ti Pb. Alloy systems of lead with metals of the transition group, such as iron, chromium, tungsten, cobalt, nickel, or manganese, generally show little liquid or solid solubility.' Craighead et al.,3 however, indicated that lead is soluble in a and ß titanium to at least 2.1 pct. They also found that approximately 50 pct of the lead added was lost during the arc-melting of Ti-Pb alloys. In the present investigation, are-melting also was used to prepare Ti-Pb alloys and similarly large lead losses were encountered. Experimental Procedure Alloy Preparation: Considerable difficulty was encountered in the melting of the Ti-Pb alloys and a number of experiments were performed before the method described was developed. Alternate layers of lead and titanium were wrapped with titanium sheet and compacted. This "sandwich" was annealed at 300" to 350 °C and was melted according to the procedure described by Cadoff and Nielsen.' Using a low current, 150 amp, and an argon atmosphere, four to eight 10 to 20 sec melts were made on each alloy, with the button being turned over between each melt to insure homogeneity. In the high lead range, the preliminary sintering was found to be of doubtful value and was eliminated. In the region above the nominal composition of 70 wt pct Pb, the alloys prepared in this manner always consolidated into two phases. The outside appeared to be composed entirely of lead and the inside was a Ti-Pb alloy. In order to determine whether this phenomenon was caused by a miscibility gap, master alloys with a nominal composition of 40 wt pct Pb were machined and mixed with various percentages of lead sheet. This mixture was compacted and melted in the same manner as the other alloys. Although these alloys were not homogeneous, they did not show the layering that was observed

Jan 1, 1956

By A. J. Shaler, J. Prieto Isaza, J. Wulff

Tantalum metal has a number of unique properties which give it widespread application in modern technology and in research. In electronic apparatus involwing high temperatures in vacuo some of the refractory metals 6ecome brittle. Tantalum retains its ductility at elevated temperatures, and this property, together with its low vapor pressure and its ability to absorb gases, make it important in this field. Its chemical inertness in the presence of acids, halogens, and other corrosive media, makes it useful in chemical apparatus and in neuro and orthopedic surgical applications. As an addition to tungsten carbide in steel-cutting grades of cemented-carbide tools, the carbides of tantalum and of its sister element colum-bium have found considerable use. The present-day method1 of manufacturing tantalum is to sinter a powder compact of the metal in vacuo. Because tantalum-columbium alloys are difficult to work and since tantalum ores almost always contain columbium, the two metals must be separated chemically, the products of the separation being generally in the form of potassium tantalum fluoride and potassium columbium oxyfluoride. These salts are used in fused-salt electrolytic baths from which coarse metallic powder is deposited as a sponge at the cathode. Finer powder can be made from the coarse product or from tantalum scrap by first submitting it to an embrittling heat treatment in hydrogen and then crushing it. The fine hydrogen-containing powder may subsequently be degassed in vacuo at a high temperature. Prompted by the success of others2,3 in the magnesium reduction of titanium and zirconium halides; one of the authors undertook4 an investigation of the thermochemistry of the reaction between solid magnesium and gaseous tantalum penta-chloride, a subject which had not previously been studied. He found that the heat of formation of tantalum pentachloride is approximately 364,250 cal per mol at temperatures lower than the melting point of magnesium. A value of 83 units was found for the standard entropy at these temperatures. With these values, it was shown that magnesium and tantalum pentachloride are capable of reacting exo-thermically at temperatures less than the melting point of magnesium, and that no great violence is to be expected when these substances are heated together. The exothermic nature of the reaction indicated that the Drocess should be efficient and should not require much externally applied energy. Tantalum pentachloride is not commonly made commercially, and any process using it therefore involves the subsidiary problem of producing and handling this substance. A series of experiments was carried out in which dried chlorine saturated with carbon tetrachloride was passed over heated tanta-

Jan 1, 1949

By Charles H. Jacoby

In 1774 a Swedish chemist, C. W. Schule, first recognized manganese as an element. That same year Schule's associate, J. G. Gahn, isolated the element manganese for the first time. In 1856 the Bessemer process of steelmaking gave birth to the economic importance of manganese. Later, in 1882, Robert Hadfield discovered high manganese steels. Although the primary use of manganese is in the ferroalloy industry, two additional important uses for manganese are making chemicals and dry cell batteries. Manganese is also vital to plant and animal life. Various chemical compounds of manganese are used in fertilizers, feeds, glass manufacture, paints, varnishes, and for numerous medicinal and chemical purposes. For the study of industrial minerals and rocks, only the chemical and battery grade material is of concern, but for the sake of comparison, metallurgical aspects are included in this presentation. Although currently the world's largest consumer of manganese, the United States is producing only minor amounts of ore at the present time. This is due primarily to the fact that although huge deposits of manganese occur within our borders, all are of such low grade as to be economically unfeasible to mine. For many years various government agencies and segments of private industry have devoted sizable sums of money and time in an attempt to upgrade ores from the various deposits, and also in an attempt to recover manganese from blast furnace slags. In view of current world prices, and the quantity of high grade material available, there is little likelihood that any of these approaches will be found to be economically feasible in the near future. Geology and Mineralogy Mineralogy There are over one hundred minerals that contain manganese. These minerals vary from those with compositions that are predominantly manganese to those having only minor percentages. [ ] Geology From a geological standpoint most manga¬nese deposits are complex. In general, it can be said that all primary deposits of manganese are carbonates or silicates. The most productive and profitable have been those of sedimentary origin and residual concentrations. Examples of these residual types of deposits are associated with the metamorphosed lodes of Madhya, Pradesh, India, and the lateritic deposits of Orissa and Bihar, India; nodules in the residual clays of the Philippine Islands, U.S. southern Appalachians and Arkansas, Ghana, and Brazil. Sedimentary manganese deposits are best exemplified by the Nikopol district of Russia and the Tchiatouri area of Georgia, USSR. Hydrothermal replacements are characterized by the rhodochrosite deposits of Butte and Phillipsburg, Mont. In general, the hydrothermal type deposit has not resulted in any large tonnages being produced, but may have given

Jan 1, 1975

By A. H. Daane, B. J. Beaudry

The yttl-irtm-nickel system has been investigated by metallo-graphic, thermal ard X-ray methods. There are nine intermetal-lic compounds present: Y3Ni, Y3Ni,, YNi,, YNi,, Y,Ni,, YNi,, and Y,Nil, which rodergo peritectic decomposition and YNi and YNi, which melt congruently. There are eutectics at 34.8, 57.5 and 93.3 at pct Ni which melt at 805°, 950 and 1285 C, respectively. Crystallographic data are given for YNi, YNi,, YNi,, YNi5and Y,Ni,. The terminal solid solubility is low. In recent work by Fisher and Fullhart' in which yttrium crucibles to be used for containers of uranium-chromium eutectic were clad with a high nickel steel, a low melting phase formed where the yttrium was in contact with the steel. To investigate the cause of this phenomenon, Daane, et al. made a survey study of yttrium systems with chromium, manganese, iron, and nickel and found an 18 wt pct Ni (weight percent) alloy to melt at approximately 950°C. voge13 in a study of the systems of cerium, lanthanum, and praseodymium with nickel found six compounds of the same formula in each system, i.e., R3Ni, RNi, RNiz, RNi3, RNi4, and RNi5. Since there is a great similarity between yttrium and the rare earths, compound formation was likely to occur also in the yttrium-nickel system. voge13 observed that the ability of transition metals to form compounds with the rare earths diminishes in the order nickel, cobalt, and iron while no compounds are formed with manganese, chromium, and titanium. Available data on yttrium systems with these elements indicate similar behavior, except that a slightly greater tendency toward compound formation is apparent. Simple eutectics with limited solid solubility have been found in the yttrium-titanium4 and yttrium-chromium5 systems, but at least one compound is present in the yttrium-manganese system.' In the case of yttrium-iron, at least three compounds are present,5 while only two are present in the cerium-iron,' and no compounds are present in the lanthanum-iron5 system. On the basis of both the trend in the alloying behavior of yttrium and rare-earth metals with the first transition series metals and the large number of compounds in the cerium-nickel system, one might expect six or more compounds to form in the nickel-yttrium system. A consideration of Hume-Rothery's rules of alloying based on size factor, electronegati-vity, and valence predicted low terminal solid solubility and possible compound formation. The present study was undertaken to confirm these predictions of

Jan 1, 1961

By K. L. Clark, Nicholas Kowall

MAXIMUM furnace efficiency and close control of final steel composition demand that the steel melter be able to follow closely the variations in the carbon content of the bath. For many years, the fracture test has been used by melters for estimating carbon, because it requires only a very short time to pour a sample into a chill mold, cool it in water, break it with a sledge hammer, and observe the fracture. After long experience gained from the observation of such fractures, most melters are able to estimate the carbon content of steel with amazing accuracy, particularly in the lower carbon ranges. Notwithstanding the utility of the fracture test, there has been widespread interest in other methods for rapid carbon determination to supplement it and to ensure against possible errors in human judgment. Several methods, based upon measurement of magnetic properties as functions of carbon content, have been described in the literature1-6 and have been found to give reasonably reliable results in plant practice when scrap composition is known and controlled. A difficulty encountered in the use of analyzers of the magnetic type is that alloying elements do change the relationship between carbon content and magnetic properties and it becomes necessary to establish calibration curves for all the expected combinations of alloy residuals that are not oxidized in the bath. The work reported here is of an investigation of a method that is independent of alloying elements normally present in constructional steels, being based upon the relatively well-known fact that in plain carbon and low-alloy steels martensitic hardness is a function of carbon content alone. WORK OF OTHER INVESTIGATORS In 1938, Burns, Moore and Archer published a curve7 that showed that the hardness of martensite is a function of carbon from 0.10 to 0.50 per cent and is independent of grain size or alloy content, at least within the range found in constructional S.A.E. compositions. Some of the deeper hardening types investigated, however, contained carbon to a maximum of only 0.20 per cent. Using this principle, Kern8 developed a method of rapid carbon determination for S.A.E. 43xx steels. His procedure consisted of chill-casting a test bar that was large enough to give a surface that could be prepared for Brinell indentations. The test piece was quenched immediately after solidification was complete and hardness readings were taken, so that the carbon content could be read from an experimentally determined curve giving the relationship between as-quenched hardness and carbon. EXPERIMENTAL WORK Although Kern's work had been done with a relatively deep-hardening steel, it

Jan 1, 1944

By W. F. Hosford

A quantitative explanation is offered for the peculiar curled grain shapes found in the microstructures of drawn wires of bcc metals and compressed aluminum specimens. It is shown that once an [011] fiber texture is developed, the slip system are so oriented that further slip tends to produce plane strain rather than axially symmetric flow of individual grains. With plane -strain flow, however, compatibility of neighboring grains can be maintained only by a bending of grains about one another. PROBABLY the most obvious microstructural feature of heavily worked metals is the distorted shapes of grains. In the course of deformation, the shape change of each grain must be compatible with that of neighboring grains; otherwise holes would be formed at grain boundaries. In analyses of poly-crystalline deformation by Taylor and subsequent workers it was assumed that grain boundary compatability was maintained by each grain undergoing the same shape change as the polycrystalline aggregate in which it is imbedded. According to this assumption, when simple tension or compression is applied to a random polycrystal or to one with a fiber texture, each grain ought to deform with axially symmetric flow. Initially equiaxed grains should become cigar-shaped during wire drawing and pancake-shaped during compression. While microstructural observations of deformed metals are generally in accord with such a picture, certain striking exceptions may be noted. Peck and Thomas5'8 found that grains in heavily drawn tungsten, iron, and niobium (columbium) wires become shaped like ribbons curled about the wire axis, Fig. 1. At a suggestion from the author, they rationalized the development of these micro-structures by considering the orientation of the slip systems in the [011] texture formed by the wire drawing. It is shown in this paper that marked departure from axially symmetric flow of grains is also observed when polycrystalline aluminum is severely compressed. Again the explanation must lie in the orientation of the slip system in the [011] texture, which is formed in fcc metals by compression.' THEORY A simple quantitative argument will be presented to show that the development of these microstructures is a result of flow on the combination of slip systems for which the minimum amount of mechanical work is expended in producing a unit tensile (or compressive) strain. First it must be realized that the macroscopic strains of a crystal resulting from slip are given by: where d dy, and dc, are normal strain increments parallel to the x, y, and z reference axes; are crystallographic shear-strain increments on slip systems

Jan 1, 1964

With only one certain exception coal was never used by the Indians, before white men came to America, for any purpose except as an ornament or for paint. Within the past few years it has been discovered that the Hopi Indians in Navajo County, Arizona, used lignite for burning pottery, about 1000 A. D., as dated by tree rings. They were able to obtain a clear color in their ware that could not be obtained with brushwood fires. The coal was mined by hammers from outcrop openings, some of which can still be seen, and the tools used have been found in them.1 (See also p. 376.) Small figurines made of splint or cannel coal have been found at various localities in western New York, in Ohio and possibly other places, evidently made for ornaments. There is no reason to think they are of any greater age, if as old, as that mentioned above. Some crushed coal was used by the Indians as a black pigment in painting themselves, but apparently the eastern Indians never used coal for any other purposes. Some think that several tribes of Indians in the Northwest, along the Missouri, used coal as fuel, the basis being a statement attributed to Radisson about 1660. In the Jesuit Relations of 1659-1660 it was stated that Radisson anti Groseilliers had just returned from a visit to Lake Superior and beyond. "In the country of the Poualak, which means 'warriors' they say `as wood is scanty in supply and small in size in their country, nature has taught them to make fire with coal from the earth (charbon de terre) and to cover their cabins with skins.'"2 This statement was only hearsay with the good fathers, and the reputation for veracity of the two explorers is not of the best, and it is extremely probable that instead of coal they used turf, dry grass or buffalo chips. In fact in Radisson's Voyages' he does not mention coal at all in his account of the third voyage, when he was supposed to have seen it used. His accounts of his travels are badly mixed, and he and his companions were much better travellers than narrators. None of the early travellers up the Missouri River saw any evidence even then that the Indians were familiar with the use of coal. In the writer's opinion it is unlikely that coal was used as stated by the Jesuits.

Jan 1, 1942

By B. M. Larsen, T. E. Brower

THIS discussion is based on statistical manipulation and evaluation of operating data from several commercial blast furnaces which include rather wide variations in practice. We are concerned here mainly with the effect of manganese on sulphur elimination under operating conditions in commercial blast furnaces, but this effect can be outlined more clearly by including certain other variables such as silicon in iron and slag basicity, which are all in part interrelated. Real progress in understanding all of the smelting-zone reactions in this process requires more laboratory studies of slag-metal melts similar to those in the blast furnace, with careful control of oxygen pressure. Such experimental work is very difficult, however, and in the meantime, with the present scarcity of more fundamental data, such statistical trends and correlations as can be obtained from commercial furnace operating records may help in choosing laboratory test conditions as well as adding to our knowledge of operating procedures. We do not know just why the furnace data are apparently so irregular that statistical treatment of large numbers is needed to obtain accurate trends. It is partly caused, no doubt, by the many simultaneous variables present, but there are perhaps also such factors as: (I) the con- tinual swings up and down so characteristic of the process; (2) difficulty of accurate sampling of co-existing phases, as well as (3) the presence of both equilibrium and rate factors as a result of the fast driving of the process. In desulphurization, for example, the flow of thin layers of slag and metal over solid coke surfaces, plus the contact between droplets of iron falling through the slag layer in the hearth, would lead one to expect a close approach to equilibrium. On the other hand, most of the critical part of sulphur removal may occur in a short time and distance below the tuyeres with perhaps an accompanying decrease in oxygen pressure and the concurrent reduction of much of the silicon and manganese, so its degree of approach to equilibrium may be quite variable. Probable Chemistry of Desulphurization Since there is little evidence that sulphur atoms in the iron solution are aasociated to any appreciable extent in molecules such as MnS, the transfer from slag to metal and back again can be written:* (FeS) ? [FeS], [I] hence the following reactions seem most probable in the slag phase, (FeS) + (CaO) ? (CaS) + (FeO), [2] and, (FeS) + (MnO) ? (MnS) + (FeO). 131

Jan 1, 1949

By H. M. Chance, T. M. Chance

The term 'Llow-sulfur coal," as used in this discussion, is limited to coals containing less, or very little more, than 1 per cent. sulfur. For certain purposes it might be advantageous to include coals that contain as much as 1.25 per cent. sulfur, because such coals have been considered as available for making metallurgic coke of fair grade; but the same reasoning might apply to coals containing as much as.1 1/2 per cent. sulfur, if facilities exist for utilizing such coal by admixture with coal containing less sulfur. , In Pennsylvania, it is not possible to define the exact geographic limits of the areas in which coals relatively low in sulfur exist, because variations in the percentage of sulfur occur irregularly and often abruptly. It is not uncommon to find areas, more or less circumscribed in extent, in which the coal is relatively high in sulfur although all the surrounding territory contains low-sulfur coal; and, similarly, small areas of low-sul. fur coal exist in districts where the coal is normally high in sulfur. These generalizations apply, with few exceptions, to the whole area underlaid with bituminous coal in the State of Pennsylvania, and also to each of the individual beds of coal in this state. Pennsylvania is poorly supplied with coals of the low-sulfur class. We find, however, as an offset to this condition, very large areas in which coal ranging from 1 to 2 1/2 per cent. sulfur exists in beds of good workable thickness. This coal, by washing, can be reduced in sulfur low enough to make coke of a high grade. While a considerable tonnage of workable low-sulfur coal remains in the Pennsylvania bituminous districts, it has been necessary to turn to the areas of higher sulfur coal for the supply of metallurgical coke. This utilization of the higher sulfur areas has been made possible either by washing or by selective mining. Washing.—Fine crushing of the entire feed is not generally necessary to the successful washing of Pennsylvania coals. This is due to the fact that the pyrite is rarely finely and uniformly disseminated throughout the coal but is usually confined to high-sulfur benches in the bed or to concentrations on the cleavage planes of the coal. Hence, by graded crushing and sizing it is frequently possible to at once reclaim those portions of the coal bed containing coal that is relatively free from pyrite.

Jan 1, 1920