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By W. M. Weigel

The rapid advance, during recent years, in the manufacture of articles that have been in common use for generations and the development of new materials entering into appliances and devices unheard of fifty years ago have created a demand for some of the more common non-metallic minerals. To supply this demand there has quietly developed a branch of the mining industry of which little is known by most mining engineers —the mineral-filler industry. In this paper, and generally in the trade, the term mineral filler is applied to a mineral used in its natural state after mechanical preparation only (such as grinding, milling and washing), without undergoing any chemical change and without exerting any chemical action on the substance with which it is incorporated. The consumption of mineral fillers is confined chiefly to large manufacturing centers in the Eastern and Northern Statcs. Where available, materials originating near the place of consumption are used, as most fillers bring a comparatively low price and, frequently, the freight may equal the cost of the finished filler at the mine. The Southern States have abundant deposits of non-metallic minerals suitable for fillers, which supply the manufacturing industries in adjacent states. Many of these deposits are equal to, or superior, in quality to those mined in the North, but they can not compete with them because of the freight rates. The mineral-filler industry has felt the present depression, but it has suffered less than most mining industries, for as a rule the mills are simple, the overhead expenses are low, and the labor is entirely local and can revert to farm work when the plant closes down. Many other deposits in the southern district will undoubtedly be developed when business revives and freight rates are adjusted. Most of the non-metallic mines are small and the tonnage handled is low in comparison with the tonnage involved in modern large-scale mining of coal, iron, copper, etc. On the other hand, at nearly all non-metallic

Jan 1, 1923

By G. E. Murray

Lead blast-furnace work at Trail is unique in that this is the only plant in the world where blast-furnace smelting on a large scale for both lead and zinc recovery is carried on under one management. On this account the factors that determine economic practice differ from those of other lead smelters. In addition to this different metallurgical outlook, a great deal of structural investigation has been carried on, which has resulted in the development of a furnace which should prove of interest to all who are faced with similar problems. This description may be conveniently divided into various heads as follows: (1) General—make-up of charge, etc.; (2) structural—(a) hoods and charging methods, (b) shaft construction, (c) crucible construction; (3) handling slag; (4) handling bullion; (5) metallurgy. The Charge The Trail plant is an old one, on a congested site. As with most old plants, this limits the method of charge preparation. In general, lead metallurgists will agree that sinter when it reaches the blast furnace should contain as small an amount as possible of actual fines. There would be some difference of opinion as to the maximum size as well as of the minimum required, but to obtain the optimum condition some form of screening is indicated. This screening should take place immediately before the blast furnace if possible, and because sinter deteriorates in storage there should be as little as possible. There should also be the minimum amount of handling between the sinter machines and the furnaces. Unfortunately, at Trail the sinter must be handled considerably before it is charged to the furnace, and no feasible scheme of screening has yet been devised. Trail practice, then, demands that the sinter be a strong one physically. Dependent upon the amount of lead in it, it has been found advantageous to incorporate granulated slag into the final sinter, the amount varying inversely as the lead content. The addition of this slag makes a harder sinter, which arrives at the blast furnace in a better condition, but it entails some lack of efficiency, because desulphuriza-tion is not so rapid nor so complete.

Jan 1, 1937

By Stanley Arnot

THE Mother Lode district in California is probably better known as the land of Bret Harte and Mark Twain than as a gold-producing district, although in this respect it holds an important place. The history of . quartz mining on the Mother Lode dates back to the '50s, when, after the first fever of the gold rush, men began to look for something more permanent, if not more lucrative, than the placers. Since that time,: the discoveries, shutdowns, and reopenings of the many mines, together, with the varying fortunes of their numerous operators, constitute an: interesting chapter in California history, but one too long to include in a description of mining methods. The district is well situated, so far as physical conditions are concerned. The Mother Lode extends along the western slope of the Sierra . Nevada Mountains, about half way between the crest and the Sacramento and San Joaquin Valleys, the country traversed consisting of low rolling hills-typical foot-hills. The elevation varies from 1000 to 1800 ft. and the climate is most desirable. There is an abundance of water, timber is easily obtained from the higher foot-hills a few miles east, and: cheap hydro-electric power is generated practically on the Lode. These natural conditions are conducive to low costs in mining and have been the deciding factor in allowing profitable exploitation of the comparatively low-grade orebodies. Transportation facilities are, not bad. Five branch railroads, from the main trunk lines in the valley, tap the district at intervals of from 10 to 30 miles, the roads connecting these are fairly good for trucking purposes. Labor is made up of several nationalities, the principal ones being American, Spanish, Jugo-Slav, Italian, and Mexican. There were few Mexicans on the Lode prior to 1919, but the shortage of labor, experienced generally by the mining camps of the country in the years following the war, made necessary the hiring of some. The percentage of Mexican labor rapidly increased in 1920 and for a year or so thereafter, but it has now diminished until at most of the mines only a few Mexicans are working. Of the different nationalities, Americans are the most efficient and Mexicans the least.

Jan 9, 1925

By Frederic H. Lahee

After abandoning two dry holes, on the Mrs. Daisy Bradford land, C. M. Joiner finally completed his No. 3 on Sept. 8, 1930, at a total depth of 3592 ft. This well is 735 miles somewhat north of west of the town of Henderson, in Rusk County. Between Oct. 2 and Oct. 6, this well made numerous heads which indicated that it would produce about 300 bbl. daily. By Nov. 27, two dry holes,' completed near the Joiner well, taken in conjunction with the small production from this well, served to dampen any enthusiasm among those who hoped for a major pool; but by December 18 three other producers had been completed. On Dec. 28, 1930, E. W. Bateman's Lois Della Crim No. 1 drilled itself in with an estimated daily yield of between 10,000 and 15,000 bbl. from a depth of 3650 ft. This well, 10 miles north of the Joiner well, also in Rusk County, initiated the leasing campaign that led to the later orgy of drilling. On Jan. 25, 1931, Moncrief, Farrell and others completed their J. K. Lathrop 1 as another big producer, 7 miles east of Gladewater in Gregg County and 14 miles N. 20" E. from the Bateman Crim well. The Lathrop well, 3574 ft. deep, flowed 320 bbl. in 1 hr. through a ½in. choke. The first well completed in Smith County was Cook No. 1, drilled by Guy Lewis et al. in the James Jordan Survey, 4.7 miles west of the Joiner discovery well. On Mar. 30, 1931, it flowed 137 bbl. in 33 min, through a ¾-in. choke. The top of the pay sand was found at 3672 ft. Finally, in Upshur County, on May 5, the Mudge Oil Co. et al. completed the Richardson No. 1, which flowed 274 bbl. through a 4 2/64-in. choke in 2 hr. This well is 4 miles northwest of the Lathrop producer. Development It is unnecessary to give a lengthy description of the rapid development of the area discovered in these five far-separated wells, or to discuss the speculations offered as to whether there were several pools or only one.

Jan 1, 1932

By Augustus J. Bowie

THE discovery of the auriferous cement gravel deposits in and near Deadwood Gulch, Lawrence County, Dakota Territory, in 1876, created a mining excitement, and rush to the Black Hills. Numerous ten and twenty-stamp mills were built on Deadwood Gulch ; Central City and the town of Deadwood sprang into existence, and became the centres of the mining industry of the Hills. Gold quartz deposits were subsequently discovered and located on a so-called mineral belt, extending from Whitewood Creek on the south, to Deadwood Gulch on the north. A northerly group of these quartz locations was purchased in December, 1877, by parties in San Francisco,, who formed a company, incorporated under the name of the Father de Smet Consolidated Gold Mining Company. This company immediately commenced the work of exploration. The deposits were proved to be extensive, and easily mined, but it was found that the ore, on an average, would not. yield more than $10.00 per ton* of 2000 pounds. At this time the cost of supplies and materials in the Hills was high. Miners' wages were $3.50 per diem, and custom mills charged $6:00. per ton. The successful working of mines on "the belt" necessitated cheap milling. The ores being of so low grade,† this could be accomplished only with large and economically running mills. To insure the greatest possible economy, the Father,& Smet mill * Up to December 31st, 1879; the de Smet mine had produced bullion of the value of $718,211.68 from 76,155 tons of ore, being an average value of $9.41 per ton. During the year ending December 31st, 1879, the mine produced 46,344 tons, which yielded $425,360.42, or $9.18 per ton. The total gross cost of working for the corresponding period, including deadwork, exchange, legal expenses, fete., wits $219,914.82. The yield of the mine for the year 1880 (Engineering and Mining Journal, February 5th, 1881), was $600,011, with total expenses of mining and milling, of $313,108. † The first 4083 tons extracted from the de Smet mine, were selected, yielding $62,783.97. The next 25,728 tons produced $230,067.24. The total 29,811 tons were weighed, and worked at custom mills. The yield of the first 409 tons from tine Golden Terra Mine was $11.44 per-ton.

Jan 1, 1882

By J. A. Jr. Church

I. EARLY FURNACES. COPPER blast-furnace construction in America has long recognized a general standard in the rectangular mater jacketed shaft with separate forehearth. The details, however, and especially those relating to shaft dimensions, have had to meet a great variety of conditions, and show no uniformity whatever. Selecting a few of the present day Western plants, we find furnace lengths varying from 12.5 to 87 ft., widths at tuyere level from 42 to 84 in., and heights above tuyeres from 5 ft. 4 in. to 22 ft. 4 in. ; and comparing present designs with those of the past, we meet contrasts no less striking. Unlike its transverse section, the length of a furnace has come to depend rather on the size of unit desired than on the requirements of smelting ; and if the question of length be ignored, several of the various existing designs can be grouped together on the basis of a common transverse section. This section, having a width at tuyere level of 56 in. and a height above tuyeres of 18 ft., was developed at Great Falls, and the story of its development contains some interesting features. Ground was broken for the Great Falls works in the spring of 1891. About a year later, the concentrator was able to begin operation, followed within the next few mouths by the Brueckner, reverberatory, and converter departments, and in April, 1893, by the first blast fur-

Jan 7, 1913

By Alexander Gakner

Since the end of World War II the Soviet Union has made great strides toward economic development of the country. Today it is firmly established as the world's second largest industrial power after the United States and it has committed itself before the people on both sides of the Iron Curtain to catch up with and to overtake us within a few decades. No nation could achieve such a degree of industrialization or aspire to become the world's industrial leader without a large supply of many basic mineral raw materials or at least easy access to such resources. In addition, it requires technical competence, a large skilled labor force, and a high rate of capital investment. Why should we concern ourselves with these Soviet aspirations? The Free World's yardstick of measuring a nation's well-being is the standard of living of the people in that country. Presumably even Communist governments are committed to some, however gradual, improvement in the standard of living of their people. The population of the USSR is 20 pct greater than that of the United States. Hence, if they ever achieve a standard of living equal to ours, they must by all logic reach a greater overall industrial output. However, in the U.S.S.R. the standard of living of the masses has not kept pace with the expanding industrial production and is still far behind that of the United States. Our concern, therefore, is with what the USSR intends to accomplish with its expanding industrial power: Will they produce large mineral raw-material surpluses and manufacture products to compete with us for world markets and attempt to put us out of business if not use it directly to wage war; or will they use their resources merely to produce more consumer goods for the Soviet people; or will they, along with other Communist nations in the not-too-distant future, compete with us for the world's sources of mineral raw-materials to maintain a high level of mass consumption?

Jan 6, 1960

By Henry S. Mackay

The interesting paper recently submitted by Messrs. Tobelmann and Potter1 shows that chemical problems have developed which are of great interest in this new and important branch of metallurgy. Those who realize what a process of this kind really involves will agree that it deserves a better name than '(leaching process." The word leaching conveys the idea that the copper is simply leached out of the ore with acid and then precipitated on iron; the chemistry of such a process would be of little interest. As, however, during the first year's operation of this plant 24,400,532 lb. (11,067,837 kg.) or about two-thirds of the produce, was electrolytic copper, while 7,992,033 lb., or about one-third, was cement copper, it would seem more appropriate to term the process electrochemical. It further appears, from the particulars given, that efforts have been made to have the entire production in the form of electrolytic copper. This could be done if the chemistry of the problem were kept in balance, so that it would not be necessary to discard any solutions that carried copper and acid of commercial value. A process, such as that under consideration, in which the copper is precipitated from solution by electrolysis, the solvent being at the same time regenerated, is one that requires special chemical and electrochemical knowledge and experience if all the copper in solution that has been extracted from the ore is to be successfully and economically produced as electrolytic copper. The object of this paper is to discuss, from the practical and theoretical standpoints, the chemical and electrochemical problems involved in the working of a process of this kind, and which must be solved to keep it in chemical and electrochemical equilibrium. In such a process, if the chemical balance is maintained, a solution can be used alternately as a solvent for the copper in the ore and as an electrolyte of high efficiency for the precipitation of the copper. To maintain the chemical equilibrium and so attain this result, the process must be so controlled that the solu-

Jan 1, 1921

By Robert B. Hall, Charles W. Bauer

Alunite, KA13(SO4)2(OH)6, was used from the 15th century until early in the present century as a source of potash alum and aluminum sulfate. Since early in this century its use for this purpose has declined to virtually nil with acid treatment of bauxite or clay replacing alunite as a source of alum. During the First World War alunite served as an emergency source of potassium sulfate fertilizer in the United States and Australia. More recently, alunite has been investigated as a nonbauxite ore of aluminum, with potassium sulfate and sulfuric acid as recoverable byproducts. The Soviet Union established a commercial-scale plant in the mid-1960s producing cell-grade alumina from alunitic ore and recovering byproduct potassium sulfate and sulfuric acid. Interest in nonbauxite ores, including alunite, is expected to continue in countries that presently are dependent on imports of bauxite or alumina to support their aluminum industries. The US has the largest aluminum producing and fabricating industry in the world but imports more than 90% of its aluminum-bearing raw materials. This chapter reviews the status of alunite as a potential source of aluminum with byproduct fertilizer, and the resources available to support this industry, whenever this becomes economically feasible. Historical Background Alunite has received little attention until recently, yet for centuries it held a higher place than many of the industrial minerals considered essential today. The deposits of alunite at Tolfa, northwest of Rome, were mined almost continuously from 1462 until about the middle of the present century (DeLaunay, 1907; Lombardi, 1977). The potash alum made from the alunite, called Roman alum, was a valued commodity in international trade, being used in the making of paper, in tanning leather, and as a mordant in textile dyeing. The industry at Tolfa is said to have been controlled by the Vatican and was a source of revenue for the Papal State until fairly recent times (Lombardi et al., 1977). Since early in the present century, the preferred method for making alum has been by acid treatment of bauxite, clay, or other aluminous material, so that alunite has fallen into near-total disuse as a source of alum chemicals. Appreciable quantities of mixed alunite-kaolin rock still are quarried in the Tolfa district for use by the Italian cement industry (Lombardi and Mattias, 1979), but the role of this material in the making of Italian cement is not clear. A large deposit of alunite-bearing rock in Chekiang province, mainland China, is said to have yielded a "great quantity" of alum over a period of several centuries (Yih, 1931). During the 19th century, alunite deposits in France, Spain, Australia, and other countries were exploited as raw material for making potash alum and aluminum sulfate. Interest in alunite was revived during the First World War, not for alum, but as an emergency source of potassium sulfate fertilizer. Large veins of nearly pure alunite had just been discovered on Alunite Ridge, 11 km southwest of Marysvale in Piute County, Utah (Butler and Gale, 1912), and these were quickly brought into production for making potash fertilizer to replace supplies cut off from the famous deposits at Stassfurt, Germany. An estimated 225 kt of Marysvale alunite was mined and processed at that time (Callaghan, 1973). Similarly, Australian alunite deposits were mined as a source of potash fertilizer during both world wars (Hall, 1978).

Jan 1, 1983

By Hugh D. Pallister, Richard W. Smith

The Alabama flake-graphite industry has flourished only in times of war when importations of foreign graphite for crucible use have been greatly curtailed or cut off. World War I was a boom period and in 1918 some 30 plants produced 7,795,475 Ib. of flake graphite. The industry languished following the war and no graphite was produced from 1929 to 1939. The Federal Bureau of Mines began investigating the Alabama graphite deposits in 1940 and worked out modern recovery methods that were applied to production during part of World War 11. A prospecting program disclosed reserves of over 25 million tons of graphtic schists. In 1942 the War Production Board authorized the enlargement of the only plant then producing flake graphite and the construction of two new plants. Production of flake graphite from these three plants in 1943 is estimated at 8,100,000 lb. Increased imports from Madagascar late in 1943 shut down one of the new plants and curtailed production at the other. Alabama flake graphite has not yet been able to compete with Madagascar graphite for use in the manufacture of any but small graphite crucibles. The future of the industry must depend on meeting the higher carbon content demanded for other uses. Introduction The flake-graphite deposits of Alabama are confined to a narrow belt extending for about 60 miles southwest from the northeast corner of Clay County across Coosa County to the edge of the Coastal Plain in Chilton County (Fig. I). The continuity of the beds is broken by a 10-mile gap between Millerville in Clay County and Goodwater in Coosa County. The Clay County beds extend about 27 miles northwest of this gap, with a width of 3 to 4 miles near Millerville and of less than a mile at the northeast end. The Coosa-Chilton are? is about 33 miles long with a fairly consistent width of 2 to 3 miles. Geologically the graphite belt lies near the northeast boundary of the outcrop of metamorphic rocks of the Ashland series, a complex group of intensely folded and faulted beds of pre-Cambrian age composed mostly of quartz-mica schist, in which the mica is predominantly muscovite; garnet-mica schist, in which biotite is more common; and hornblende schist. Flake graphite is found in commercial quantities only in the quartz-mica schist. The best graphite, both in quantity and quality, is found in schist cut by numerous small pegmatite stringers along the planes of the schistosity. The pegmatites themselves contain little graphite, but the adjacent country rock is normally rich. . The schist was derived through the recrystallization of shale during the intrusion of the pegmatites, the graphite being derived from amorphous carbon in the shale. The intrusion of the pegmatites, and therefore the size and quantity .of the graphite flakes, seems to have a structural control.

Jan 1, 1948

By Hugh D. Pallister, Richard W. Smith

The Alabama flake-graphite industry has flourished only in times of war when importations of foreign graphite for crucible use have been greatly curtailed or cut off. World War I was a boom period and in 1918 some 30 plants produced 7,795,475 Ib. of flake graphite. The industry languished following the war and no graphite was produced from 1929 to 1939. The Federal Bureau of Mines began investigating the Alabama graphite deposits in 1940 and worked out modern recovery methods that were applied to production during part of World War 11. A prospecting program disclosed reserves of over 25 million tons of graphtic schists. In 1942 the War Production Board authorized the enlargement of the only plant then producing flake graphite and the construction of two new plants. Production of flake graphite from these three plants in 1943 is estimated at 8,100,000 lb. Increased imports from Madagascar late in 1943 shut down one of the new plants and curtailed production at the other. Alabama flake graphite has not yet been able to compete with Madagascar graphite for use in the manufacture of any but small graphite crucibles. The future of the industry must depend on meeting the higher carbon content demanded for other uses. Introduction The flake-graphite deposits of Alabama are confined to a narrow belt extending for about 60 miles southwest from the northeast corner of Clay County across Coosa County to the edge of the Coastal Plain in Chilton County (Fig. I). The continuity of the beds is broken by a 10-mile gap between Millerville in Clay County and Goodwater in Coosa County. The Clay County beds extend about 27 miles northwest of this gap, with a width of 3 to 4 miles near Millerville and of less than a mile at the northeast end. The Coosa-Chilton are? is about 33 miles long with a fairly consistent width of 2 to 3 miles. Geologically the graphite belt lies near the northeast boundary of the outcrop of metamorphic rocks of the Ashland series, a complex group of intensely folded and faulted beds of pre-Cambrian age composed mostly of quartz-mica schist, in which the mica is predominantly muscovite; garnet-mica schist, in which biotite is more common; and hornblende schist. Flake graphite is found in commercial quantities only in the quartz-mica schist. The best graphite, both in quantity and quality, is found in schist cut by numerous small pegmatite stringers along the planes of the schistosity. The pegmatites themselves contain little graphite, but the adjacent country rock is normally rich. . The schist was derived through the recrystallization of shale during the intrusion of the pegmatites, the graphite being derived from amorphous carbon in the shale. The intrusion of the pegmatites, and therefore the size and quantity .of the graphite flakes, seems to have a structural control.

Jan 1, 1948

By I. A. Lesk

The various materials utilized in the construction of integrated circuits, and the resultant materials interfaces, are discussed with emphasis on a materials system that is compatible with all types of integrated circuits and minimizes the number of different interfaces. Problems with interfaces may originate in the silicon, on its surface, and between other materials within the integrated-circuit package independent of the silicon. ThE difference between integrated-circuit theory and practice may generally be spelled out in terms of the difficulties in fabricating a desired structure. There are, of course, problems in forming the requisite materials, of the necessary composition and geometry, where needed. A distinct class of problem, however, is associated with interfaces— generally between different materials, but occasionally between regions of the same material. We require that each interfacial system perform its proper electrical, mechanical, and thermal function, and do so in an unchanging manner with time under the most severe operating conditions. In a practical sense, we deal in nonideal systems, such that it is generally not possible to form an interface exactly the way it was intended, even if we allow for a knowledge of the practical limitations of the fabrication system. Another way of looking at this situation is to realize that, when we form a group of interfaces under apparently the same experimental conditions, there are often differences of important consequence between them. Very high-magnification pictures, such as are obtained by an electron microscope, reveal fine structure that often differs over the area of a given sample, as well as from sample to sample. Low-magnification pictures look better, but don't eliminate the problem. The trend towards integrated circuits, and away from circuits fabricated utilizing separate components, is very strong in many areas of electronics. A typical silicon monolithic integrated circuit, still in place in the wafer on which it was fabricated, is shown in Fig. 1. The increased reliability which is being experienced through the use of integrated circuits can, to a large extent, be credited to the reduced complexity of the interface system that is present in the electronic system. A glance at Fig. 1, however, with all of its epitaxial and diffused p-n junctions, and interconnect system, shows that interfaces have been far from eliminated. Fig. 2, depicting a schematic cross-sectional view of a (partial) typical integrated circuit, illustrates in detail the numerous interfaces that occur in this type of structure. Besides the individual circuit elements, there are interfaces for isolation, and, Fig. 3, even a crossover requires a series of interfaces. What has been accomplished, however, irrespective of whether or not the total number of interfaces in the circuit is increased or decreased, is that the variety of interfaces is very much reduced. This is because there are fewer materials involved. Since the expense incurred in an R & D program to improve reliability is pretty much directly proportional to the variety of interfaces involved, it is economically advisable to work with integrated circuits for highest reliability. That is not to say that circuits built using standard components could not be made as reliable as integrated circuits, but that the expense involved because of the more complex over-all interface system would be much greater. Although electronic performance utilizing integrated circuits with all silicon elements (i.e., transistors, diodes, resistors, capacitors) is adequate for many applications, there are numerous instances where thin-film passive elements must

Jan 1, 1965

By L. Luini, E. Lee

An exploratory survey of the heat treatment response of commercial titanium alloys (Ti-150A, RC-130B, and MST 3AI-5Cr al-loys) shows a wide range of possible hardness and microstructural characteristics. The hardening is primarily dependent on the solid state transformation of the phase. An age hardening reaction which apparently is associated with ß decomposition and precipitation has been shown. Brittleness and notch sensitivity appear to be characteristic of age hardening to high hardness. Ductility, tensile strengths, and elongations are given. A NUMBER of commercially produced titanium-base alloys have been made available by the titanium metals industry. These alloys were developed to combine the light weight of titanium with increased strength through alloy additions. Usually the properties of these alloys are reported for material in the hot rolled and annealed condition. Since the production of commercial titanium has preceded the investigation of titanium-base alloy systems, the potential of these alloys is not well known. Although heat treatability has been indicated for these alloys, data on the response to heat treatment are quite limited. The heat treatability of titanium is based on the allotropic transformation of the high temperature, body-centered cubic ß structure to the low temperature, hexagonal close-packed a structure at approximately 1625 °F. This transformation in the pure metal is of little practical significance in heat treatment since the high temperature phase is not amenable to controlled transformation. Alloy additions serve to alter the transformation kinetics and permit control of the transformation by heat treatment variables. The effects of different alloys on the allotropic transformation in titanium classify them in two general categories,1 a stabilizing and ß stabilizing elements. The a stabilizers, represented by aluminum and oxygen, raise the temperature limits of the low temperature phase and introduce a two-phase a + ß temperature range. The ß stabilizers, represented by molybdenum and columbium, lower the equilibrium temperature of ß and give rise to an extended temperature range where both a and ß phases are stable. Another type of ß stabilizer, represented by iron, chromium, and manganese, introduces a eutectoid reaction. Some commercial titanium alloys belong to complex multiple element systems. Based upon the phase diagrams of the binary systems, the general effect of commercial compositions is to vary the stability of the high temperature phase and significantly alter its transformation characteristics. Therefore it is expected that the characteristics of alloyed titanium can be altered by heat treatment. The present investigation was undertaken to survey the variations in hardness, microstructure, and mechanical properties as affected by heat treatment variables. Commercial Alloys lnvestigated The commercially available titanium-base alloys selected for this investigation and obtained as 3/4 in. diameter hot rolled and annealed bars are listed in Table I. The producer did not determine the chemical composition of the Ti-150A heats X1052 and L685. The early heats of Ti-150A, prefixed with the letter X, were melted with tungsten-tipped electrodes and cast in approximately 500 lb ingots 12 in. in diameter. The heats prefixed with the letter L were melted under a revised melting practice yielding approximately 1300 lb ingots 16 in. in diameter. The final annealing treatment was at 1200" to 1250°F for 1 hr. The MST-3A1-5Cr alloy was melted in an induction furnace and cast in 7 in. diameter ingots. This practice produces high carbon material and some variation in the amount of carbides was found in different sections of the same bar. The RC-130B alloy was melted with a carbon arc. This alloy showed approximately 2 pct carbides. No large segregation effects were observed.

Jan 1, 1955

By Bailey Willis, George Otis Smith

I. Summary of General Features. Location.—The Clealum iron-ore deposits occur on Clealum river, a tributary of the Yakima river; in the eastern spurs of the Cascade range, Washington.† Mount Stuart, 9470 feet in altitude, is a dominant peak of the district, which will therefore be referred to as the Mount Stuart district. By way of the valley and cañon of the Clealum, the ores are 18 miles from Roslyn, the colliery-town, which is reached by a short branch from the Northern Pacific Railroad. Where they occur the valley has an elevation of 2900 to 8000 feet above sea; that of the end of the branch railroad at Ronald is 2345 feet. The grade of the valley from near Ronald to the junction of the Clealum and Salmon Lasac rivers, 144 miles, is 20 feet per mile; thence to the iron-ores it is 130 feet per mile, through a cañon 241/2 miles long. Abreast of the deposits the western wall of the valley is formed by precipices of sandstone; the eastern slope is comparatively gentle, being cut in easily eroded

Jan 1, 1901

By C. R. Garr, A. R. Troiano

FLAKING or hair-line crack formation has been a major problem confronting the producer of large alloy steel forgings.' Today it is generally conceded that hydrogen in one or more forms in alloy steel sections is a requisite for flaking. Hence an obvious remedy is a thermal cycle that reduces the hydrogen content to some low level. In practice, this generally means holding the material at some subcritical temperature for a period of time depending on the size of the ingot or forging. Another method that will suffice in some cases is to cool large forgings slowly, i.e., bury in ashes. In practice, this hydrogen bake is sometimes accemplished inadvertently by a series

Jan 1, 1958

Covering resources, production, uses, defense impact, and future outlook of nickle, Mining Engineering presents the third in a series of articles on strategic minerals. Preceding it are "Cobalt" in January and "Sulphur" in May.

Jan 8, 1951

By Henry U. Ross

The behaviour during reduction of iron-ore pellets and sinters was investigated. Several methods for the preparation of agglomerates and their reduction were tried. Commercial agglomerates from world-wide sources as well as experimental samples were tested. The effect of coke additions to pellet and sinter plant feeds was given special consideration. An analysis of the test results has helped to explain many of the phenomena associated with the reduction of iron-ore agglomerates such as swelling, loss in strength and wear resistance, changes in porosity, changes in mineralogy and the formation of metallic iron and slag. A double-layer pellet with all of the coke concentrated in the core proved particularly interesting. Although large quantities of coke showed that partially reduced agglomerates containing metallic iron could be produced in the laboratory, the problem of producing reduced pellets and sinters by the existing agglomerating processes still remains.

Jan 1, 1977

By J. R. Campbell

During the past few years, the writer has had occasion to take several excursions into the realms of the washability of beds 3 and 4 of the Pocahontas coal and the proper handling of these coals in prepared sizes Fig. 1.—Washability curve for 4 by 2½-in. egg, PocahoKtas coal. Note that definite break occurs at about 1.60 sp. gr., and washing at this gravity removes all heavy rock and slate, amounting to about 5 per cent. Curve 2 indicates a bony coal, and washed egg coal might not have a good appcarance, and would be condemned by visual inspection, although it would be well prepared mechanically. to attain the greatest realization for the operator and yet produce a market product acceptable to the consumer. Preparation is not new in the southern field, but some new angles have presented themselves in recent years on account of the progress of mechanical preparation. Mechanical methods are entirely devoid of the five senses given to

Jan 1, 1931

SEC. 1. The membership of the 'Institute shall comprise six classes, namely: 1. Members; 2. Honorary Members; 3. Senior Members; 4. Associates; 5. Junior Members; 6, Rocky Mountain Members. All shall be equally entitled to the privileges of membership, excepting that no member of any class whose residence shall be outside of the United States of America, except those residing in Mexico and Canada, shall be entitled to vote. SEC. 2. Members: A person to be eligible for election or transfer into the class of Member must be at least 27 years of age and must have had at least six years' employment in the practice of engineering, mining, geology, metallurgy, or chemistry, during at least three years of which he must have held positions of responsibility in one or more of these fields. This shall not apply to Rocky Mountain members.

Jan 1, 1932

By N. Mungan

Formation damage, i.e.. reduclion in permeability, has been generally attribuled to clay minerals which expand or disperse upon contact with water that is less saline than the connate water. Luboratory, studies show that penneahility reduction can also occur in formalions containing only nonexpandable clays such as illite or kaolinite, and can be caused also by changes in pH. Furthermore, pH changes can damage even formations that are essentially free of clays. It is suggested that permeability reduction is due to the small passages being blocked by particles, which may be dispersed clays, cemenlion material or other fine parricles. These particles are dislodged by dispersion of clays due to changes in salinity or by dissolution of calcareous cement by acids, or of silicaceous cement by alkaline solutions. In working with reservoir cores, it was found that extracted cores damaged more easily and extensively than nonextracted cores. The extent of damage depended also on tenlperatltre. INTRODUCTION Permeability is an important property of porous media and has been the subject of many studies by engineers and geologists. Many of these studies are conccrned with formation damage, i.e., reduction in permeability, resulting from exposure of oil-producing formations to water substantially less saline than the connate water. This effect causes understandable concern since during drilling, completion and production phases formations are often exposed to fresh water. The damage resulting from contact with relatively fresh water has been attributed to expansion and dispersion oF clay minerals. During laboratory investigation of the use of NaOH as a wettability reversal agent to increase oil recovery from oil-wet reservoirs, several cores used in the displacement studies suffered loss in permeability. Despite the traditional usage of NaOH for conditioning aqueous mud systems, the role of the caustic filtrate in wellbore damage seems to have been overlooked. Browning2 as recently reported on the effects of NaOH in dispersing clay minerals but he was concerned only with complications that may arise in drilling massive shale beds. The following study was made to examine the role of pH and salinity changes in core damage. Where cores from reservoirs were used, tests were performed with extracted and nonextracted cores both at room and reser- voir temperatures, since it was felt that the test environment and core condition may affect the results. Because of its limited coverage and exploratory nature, this study is not intended to provide answers to field formation damage problems. It is hoped that it will encourage research into new aspects of the permeability reduction problems, particularly those allied to new recovery and production processes. PROCEDURE In all permeability tests, fluids were pumped through the cores at a constant volumetric rate. Only deaerated fluids and reagent grade chemicals were used. The fluids were passed through two ultrafine filters before injection to remove any entrained particles. The cores, with the exception of the unconsolidated cores, were mounted in Hassler holders. Water was used to transmit pressure to the sleeve. The inlet endpiece had two entry ports which permitted scavenging one fluid with another to avoid any mixing in the small holdup volume. The cores were flushed with CO2 gas, evacuated for 5 to 6 hours and saturated with the first liquid at a pressure of 1,000 psi for 24 hours to eliminate any free gas from the cores. Pressure differences up to 20 psi were measured by transducers, calibrated in inches of water and continuously recorded. For greater pressure drops, gauges were used. All reservoir cores were cleaned with a light refined mineral oil, then with heptane, and finally dried with CO2. Compatibility tests showed that no precipitates formed when mineral oil and the crudes were mixed. Some cores were extracted in Dean-Stark-type solvent extractors using xylene and trichloroethane and dried in a vacuum oven at 450F. Each test consisted of a sequence of water, test solution, and again water flow. RESULTS AND DISCUSSION STUDIES IN BEREA CORES Salinity Contrast Berea cores 2-in. in diameter and 12-in. long were cut from sandstone quarried in Cleveland, Ohio. The clay minerals were identified by X-ray diffraction to be chlorite, kaolinite, illite and incerlayered illite. Flow of fresh water or 30,000 ppm brinc does not cause any permeability reduction (Fig. 1). However, after injection of brine the core is readily damaged by fresh water. Damage starts almost instantly as the fresh water injection is begun, and at a cumulative injection of 1.2 PV fresh water, the permeability has dropped from 190 to 0.9 md. Upon continued injection, the effluent contains clay minerals dislodged from the core. The final core per-

Jan 1, 1966