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By E. J. Eisenach, Edward Matsen

AT the present time the Climax Molybdenum Co. is the largest molybdenum producer in the world and the operator of the largest underground mine in North America. It has grown steadily and rapidly since its beginning in 1917. Only 250,000 lb of MoS2 were produced during the first year from 1917 to 1918 when the company was formed. Operations were suspended shortly thereafter because of the lack of a market and knowledge of the uses of molybdenum. During World War I it was used as a substitute for some rare metals and for minor chemical purposes. After 5 years of intensive research and sales promotion, Climax metallurgists succeeded in finding proper applications for the newly discovered metal and the mine was reopened in 1924. Demand for molybdenum increased gradually and in 1930, 3.5 million lb of MoS2 were produced. In the meantime, research work for additional uses of molybdenum continued until it had established itself as one of the most important alloying elements in producing high quality steels. After the start of World War 11, from January 1942 to October 1953, 43 million tons of molybdenum ore were mined and 484 million lb of MoS2 were produced. Climax mine is located in Colorado, 100 miles southwest of Denver and 12 miles northeast of Leadville on the Continental Divide of the Rocky Mountains. The principal mine workings range from 11,170 to 12,000 ft in elevation. Topographically, it lies in the base of a large cirque that is surrounded by Mt. Bartlett to the north, Clinton Peak and McNamee Peak to the east and Ceresco Ridge to the south. This cirque opens to the west and to the Ten Mile Valley.

Jan 3, 1954

THE Searles Lake orebody is located in the north- west corner of San Bernardino County. It is a dry lake bed with an exposed salt surface covering an area of 12 square miles. Recoverable mineral values are contained in the mother liquor below the surface of the lake. Stratification in the lake bed has separated the brine into two bodies which differ in composition. Although liquor is processed from both bodies, this paper will discuss only the upper structure brine. Fig. 1 illustrates a typical cross-section of the two commercial orebodies. The orebody is composed of a porous salt deposit 70 to 90 ft deep. The upper structure is separated from the lower orebody by a 12 to 16-ft thick impervious mud seam, as shown in Fig. 1. These salt structures are composed of 55 pct solid-phase salts and 45 pct voids which are filled with the original mother liquor.

Jan 2, 1954

By Donald W. Scott

THE year just ended was the 50th anniversary of that first flight at Kittyhawk, N. C., in 1903. In this 50-year period the aviation industry made startling developments from flight distances of 120 ft-at snail's pace-to speeds of over 1300 mph and flights that circle the globe. And 1953 also was the 50th anniversary of the Utah Copper Co., the Father of our mass production porphyry copper ventures. How do developments in the mining industry stack up with the aviation industry? The mineral industry can be as proud of developments over the past 50 years as the aviation industry. Only for the developments of the mineral industry could aviation and other allied industries advance as they have. The year 1953 made its contribution to this progress, with the normal number of process and equipment developments and an abnormally high amount of activity in new plant capacity, particularly copper. Tonnages are on the increase and operating troubles are becoming more and more problems in materials handling and control. We're approaching molecular dissemination of ores with hydro and pyro methods coming to the front to take over where flotation leaves off.

Jan 2, 1954

By Kenneth L. Cook

GEOPHYSICAL exploration is continuing to expand in scope and variety of methods, in experimental studies of specific applications, and in development of theory for complex situations. Aerial surveys have been made economically more attractive by simultaneous recording of magnetic and radioactivity data, and aerial electromagnetic surveys have been attempted. The most significant success of mining geophysics was the discovery of huge lead-zinc-copper sulphide ore deposits in New Brunswick, Canada. Spectacular geophysical astivity has occurred in the area as a result of these discoveries, and geophysics is expected to continue to play a large role in this drift-covered area. Other developments this year find the Atomic Energy Commission making available monthly reports of the location of anomalous radioactivity detected in aerial and ground surveys and the AEC is installing supersensitive radioactivity assaying instruments in about 25 localities in the U. S. Interest in helicopters is being aroused because of potentialities for near-ground surveys and better location control of anomalies. Japan can be added to the list of countries using geophysical surveys extensively.

Jan 2, 1954

By Robert E. Thurmond, E. D. Spaulding, Walter E. Heinrichs

DISCOVERY of a commercial ore deposit has resulted from a program emphasizing geophysical techniques and concentrating upon exploration of alluvial-covered areas. United Geophysical Co. selected these areas as offering the best possibility of success, for here the surface evidence of ore deposits has been hidden from the eyes of prospector and geologist alike. Pima mining district, Pima County, Arizona, was selected for initial exploration. The project was successful in that a commercial ore deposit was discovered and proved through drilling and extensive underground development.

Jan 2, 1954

By Helen L. Cannon

BOTANICAL methods of prospecting for metalliferous ores are based on the premise that deposits at depth may be reflected chemically in surface vegetation. For the past several years, on behalf of the Atomic Energy Commission, the U. S. Geological Survey has investigated on the Colorado Plateau the application of botanical methods to uranium prospecting. Results suggest that orebodies may be found under a considerable cover of barren sandstone through the absorption of uranium and associated trace elements by plants growing at the ground surface. Plant physiologists have demonstrated that uranium is a necessary nutrient for plant growth and therefore is present in all plants.' The amount present is small but measurable. The absorption of uranium by plants growing over ore deposits forms the basis, therefore, for a method of prospecting by plant analysis.

Jan 2, 1954

By Strathmore R. B. Cooke, Iwao Iwasaki, C. S. Chang

PYRRHOTITE has long been considered a gangue mineral to be eliminated as tailing in the treatment of various sulphide ores. However, in recent years the world-wide lack of sulphur resources has called attention to this mineral as a potential source of both sulphur and iron. Its importance as an economic mineral, however, has not been particularly emphasized. For this reason very little is known about its response to flotation, except that it can be depressed easily in alkaline circuit, by long aeration,1,2 addition of oxidizing agents3 or by starch.4 The object of this work was to study the floatability of pyrrhotite. This includes the effect of oxidation by aeration, of copper activation, and of change in pH.

Jan 2, 1954

By J. Leja, J. H. Schulman

FROTH flotation is usually effected by the addition of a collector agent and a frothing agent to an aqueous suspension of suitably comminuted mineral ores. The action of collectors is to adsorb onto the surfaces of minerals to be separated, sensitizing them to bubble adherence. The action of frothers has, in the past, been accepted as that of froth formation only, brought about by a lowering of the air/water interfacial tension. Substances capable of producing froth are classed1a,b according to their relative capacities for production of froth-volume and froth stability in the simple frother-water system. The purpose of this paper is to show that the surface active agents acting as frothers become effective only when there is a suitable degree of molecular interaction taking place between collector molecules and frother molecules at the air/water and solid/ water interfaces. Further, the discussion will demonstrate that the actual mechanism of adherence of an air bubble to a suitably collector-coated particle is due to the molecular interaction collector-frother. This leads to the formation of a continuous interfacial film of associated molecules, anchored to the mineral by polar groups of the collector, and enveloping the whole bubble. The tenacity of adhesion mineral-to-bubble results from the strength and the visco-elasticity of this mixed film.

Jan 2, 1954

By S. H. Ash

SAFETY records have improved in all branches of the mineral industry. While annual production was rising from $2 billion in 1910 to nearly $12 billion in 1950, fatalities decreased from 3539 in 1911 to 974 in 1951. And, although safety progressed under hand methods, it progresses more effectively under modern, mechanized methods. But powered devices such as electric or diesel units can in themselves create new hazards unless proper safeguards are instituted. Health Use of diesels underground is increasing. Diesel equipment has been used in construction of 34 tunnels in 16 states in the last 17 years. The 77 noncoal underground mines in 24 states using diesels underground range from mines with large open stopes to those with narrow veins, so that the ventilating problems represent a good cross-section of U. S. mining. The U. S. Bureau of Mines has received no report that anyone has suffered ill effects from the proper use of properly maintained equipment. The necessity for correct maintenance and use of satisfactory exhaust conditioners cannot be emphasized too strongly.

Jan 2, 1954

By George M. Schwartz

WHEN reviewing the progress made in mining geology for the year 1953, one might say that not much has been accomplished and, indeed, in a subject such as economic geology not much progress should be expected in the 10 months of the year which elapsed before this article was prepared. This seeming lack of progress, however, disappears when the facts are investigated a bit. There undoubtedly are more economic geologists at work than ever before. It is probably not an exaggeration to say that more geologists are now employed by a single company in the Lake Superior district than were employed in the whole district 20 years ago. In most mining districts geologists are busy with numerous problems of production, development, and exploration and some give an occasional thought to the scientific problems of the science, or rather the art, of mining geology. In fact, this very intense activity of mining geologists is one reason why progress seems slow; they do not have time to write up results of investigations. There is also a great amount of work being done which is not released because of company policy. Conditions have improved very much in recent years but mining corporations, in general, have some distance to go before attaining par with the petroleum industry in this respect.

Jan 2, 1954

PRODUCTION of nonmetallic mineral raw materials used by the fertilizer, ceramic, chemical, and construction industries and many others, was greater during 1953 than in any previous year. Production of fertilizer minerals was the highest in history despite somewhat decreased consumption in the second half of the year. Marketable phosphate rock production increased 10 pct to an estimated total of 13.1 million LT, valued at $78.5 million, and potash production was estimated at 1.8 million tons of K2O equivalent, valued at over $60 million. Crude gypsum output exceeded the 1952 total and approached the record of 8.66 million tons mined in 1951, with indications that 1953 will show a new all-time high. Gains and losses among gypsum products were irregular, with major gains in plate- glass and terra-cotta plasters, sanded building plasters and miscellaneous building plasters, and principal declines in base-coat plasters, prepared finishes and tile. Lath and wallboard outputs were each about 20 pct higher than in 1952.

Jan 2, 1954

By R. L. Loofbourow, Roger V. Pierce

THE degree to which our mineral requirements are being met by mining leaner, more complex or more remote deposits is indicated by developments during 1953. Heavy capital investment, skilled technical advice, planning, and able execution are requirements for the typical venture. Producers, faced with rising costs and with softer prices, are seeking improved mining and recovery. Developments in production of iron ore are typical of the course being followed in metal and nonmetal mining. Iron ore development in the U. S. is concentrated on the spectacular work with Lake Superior taconite, while in Canada new production is expected from Labrador and output from Steep Rock, Algoma, and Newfoundland is being increased. Shipment of iron will soon be made from Venezuela.

Jan 2, 1954

By J. E. Andersen, J. Halpern, C. S. Samis

In the presence of oxygen, galena is oxidized in an aqueous medium containing sodium hydroxide, in accordance with the following reaction: PbS + 2O2 + 3OH ? HPbO2 + SO4 = + H2O A novel method was devised for following this reaction which takes place in an autoclave under oxygen pressure, by measuring the concentration of HPbO2- in the solution with a cathode ray polarograph employing stationary platinum electrodes. Using this procedure a study was made of the kinetics of the reaction, in which the effect of a number of variables, including temperature, oxygen pressure, NaOH concentration, and agitation, on the rate, were determined. The results of this study are described and discussed, in terms of possible mechanisms for the reaction. IT is known that sulphide minerals can be oxidized in aqueous media in the presence of oxygen under pressure. Reactions of this type can be classified in two categories depending upon whether the products of oxidation are insoluble or soluble in the aqueous medium. An example of the first type of reaction is the oxidation of pyrite in alkaline solutions, giving rise to an insoluble iron oxide. The kinetics of this reaction have been investigated and its mechanism established.' It has been applied in the oxidation of iron sulphides present in refractory gold ores to improve the subsequent recovery of gold by cyanida-tion.' The second type of reaction finds application where it is desired to recover the metal by leaching during oxidation. In this case a medium is selected in which the oxidized mineral is soluble. For example, nickel, copper, and cobalt sulphide ores can be treated by oxidation in the presence of a solution of ammonia dissolving the nickel, copper, and cobalt as the metal ammine sulphate salts.:' A similar treatment has also been reported to apply to zinc sulphide ores.' It is known that lead sulphate is soluble in solutions of sodium hydroxide or ammonium acetate. It should therefore be possible to dissolve the lead from galena ores by aqueous oxidation with oxygen under pressure using either of these solutions. The applicability of the ammonium acetate treatment has been investigated and confirmed in a recent study.' The present paper describes the results of a similar study in which the kinetics of the oxidation of galena in sodium hydroxide solutions have been investigated. The object of this investigation was pri- marily to obtain information of a fundamental nature relating to the kinetics and mechanism of this reaction. The use of pulps of comminuted ore in a study of this type is undesirable because of the difficulty in controlling and measuring the surface area. The measurements were therefore made using galena crystals of measured surface area and in this way absolute reaction rates were obtained. The reaction was found to proceed as follows: PbS + 2O2 + 3OH ? HPbO2 + SO4 + H2O [I] The products are sodium plumbite and sodium sulphate salts which dissolve in the aqueous solution. The reaction studies were carried out in an autoclave in which a desired pressure of oxygen was maintained. The reaction was followed by measuring the concentration of lead in the solution with a cathode ray polarograph. The results obtained in this investigation are presented and discussed in the present paper. Experimental Materials: The crystals of galena were obtained from Violamac Mines, Sandon, B. C. The following impurities were indicated by spectrographic analysis carried out by the Provincial Assay Office, Victoria, B. C.: Sn, 0.02 to 0.2 pct; Sb, 0.07 to 0.7; Zn, 0.01 to 0.1; Si, Fe, Mg, As, less than 0.05 pct; Ag, 126 oz per ton, the latter determined by standard fire-assaying procedure. After cutting a specimen measuring approximately 5x7x12 mm, the crystal surfaces were ground parallel to the 100 axes using No. 2 emery, washed with water, and the dimensions were measured with a micrometer. Oxygen gas was of commercial grade and supplied in cylinders by Canadian Liquid Air. Solutions were made up with chemicals of chemically pure grade and distilled water. Equipment: Autoclave: The autoclave used in these studies was constructed of stainless steel and

Jan 1, 1954

By J. E. Hawley

"The Aldermac copper deposit is in Beauchastel township, Abitibi county, Quebec. It lies about 11 miles west of Noranda. It has been described by Cooke, James, and Mawdsley(2) and by Bruce(3) and has been mapped in detail by W. Samuel and the late K. W. Fritzsche, former Mine Geologist, and by J . E. Gill and A. R. Byers, so that most of the particulars contained in the description that follows are largely the result of the work of others than the writer, whose connection with the deposit, in a consulting capacity, extended over the last 5 years of the life of the mine.According to A. V. Corlett, former Mine Manager, the property was staked during the early winter months of 1923 because of its location adjacent to the syenite intrusive that had been mapped by Cooke near the height-of-land in this township. An examination by Mr. Corlett in 1924 revealed mineralized zones to the north of MacKay lake, and in the vicinity of these K. W. Fritzsche, by dip needle and trenching, discovered the No. 1 and No. 2 deposits, which were both too small and too low grade for exploitation. During the winter of 1925, A. A. MacKay, with a dip needle, picked up a magnetic 'anomaly' in a drift-covered area about 600 feet to the north of previous workings. Trenching here revealed only a narrow zone of mineralization of little consequence, but, farther east, MacKay discovered a more pronounced 'anomaly' over what later drilling and underground work proved to be ore of commercial grade (No. 3 and No. 4 orebodies)."

Jan 1, 1954

By H. M. Zoerb

Based on the belief that operating details are a definite contributing factor to major economies, this paper traces the development of crushing cavity design in Symons cone crushers to attain maximum liner utilization. Wear rates are analyzed and compared in this presentation and drawings illustrate succeeding design changes. IN these times of rising labor and material costs, it has become more and more necessary that attention be paid to some operating details which, in their obscurity, may he the key to major economies. Liner wear in crushing cavities of secondary and tertiary crushers can become an appreciable cost item when the material to be crushed is hard and abrasive. This item of cost not only includes the value of the crushing members, but also more intangible costs such as labor and lost production due to more frequent replacement. The variables which are encountered in ores and minerals to be reduced; the design of plant and machine application; the sizes, shape, and fineness, characteristics of the crushed product; the moisture; hardness; friability; and abrasiveness of the material to be crushed are all influencing factors which must be taken into consideration in the selection of a crusher, and particularly in the design of crushing cavity and liners to be used in a crusher. Through a research program undertaken in cooperation with many operators of Symons cone crushers a new approach to crusher cavity design was made, resulting in the development of liners for specific operations which showed: 1—maximum utilization, as high as 70 to 80 pct of original weight of metal, and 2— maximum capacity of unit during the greater portion of its life. It has been found that liners so designed for a given operation will show added economies in power consumption, maintenance, and general wear and tear on the crushing unit. Initial work in the so-called tailoring of crushing cavitles was begun on the tertiary or fine crushing units where as a rule reduction ratios were low, varying from 3 to 6. Parallel or sizing zones in the lower portion of the crushing cavity were too long, resulting in a tendency to pack. It was found that very little additional crushing was done in the parallel zone after the initial impact in that zone and that a relatively small amount of' additional crushing was done by attrition, which required very careful feed control. A small amount of over-feeding would result in packing which not only consumed power but caused unnecessary liner wear as well. The illustrations which follow in this discussion will show only contours of crushing cavities, and for purposes of simplification the cavities will be considered only in their closed position. The first step, therefore. was to reduce the sizing zone to a minimum. This was done by removing the lower portion of the liner as shown in Fig. 1. The result of the change was a saving of 15 to 20 pct in liner cost, less power consumption, with no change in capacity. This change in design, while an improvement, did not go far enough. As wear took place, the change in the liner was not uniform throughout its entire length, resulting in a restriction of the feed opening and thereby loss of capacity. Furthermore, progressive wear of the liner had the effect of lengthening the parallel zone until finally the entire crushing cavity was all parallel zone, see Fig. 2. It is obvious from the reduced feed opening of the worn liner that the ability of the machine to receive material is lessened considerably. Furthermore, the long parallel zone with its worn, irregular profile did not operate at its highest efficiency. The first attempt to overcome this difficulty was carried out on a 5 1/2-ft crusher installed in a plant producing roofing granules. The material being crushed was a very hard graywacke and the crusher was closed-circuited with a screen having .232-in. slotted openings. A radical change in contour was developed, as illustrated in Fig. 3. Equal wear lines on both concave and mantle are designated 1, 2, 3, etc. The method of development of this contour is as follows: Since adjustment for wear is vertical, corresponding intersections of wear lines and vertical lines developed concave and mantle contours which maintained equal but lengthening wear surfaces in the parallel zone. The ideal contour, of course, is one in which the length of the parallel zone remains constant, but because of present foundry practice and heat treating characteristics this is impossible.

Jan 1, 1954

By Frederick F. Frick

SPONGE iron as produced at Anaconda is a fine, -35 mesh, impure product, about 50 pct metallic iron, obtained from the reduction of iron calcine at a temperature of 1850°F by use of coke resulting from slack coal. The metallic iron particles are bulky and spongey and precipitate copper readily and rapidly from a copper sulphate solution. Investigation of the treatment of Greater Butte Project, Kelley, ore at Anaconda early showed the desirability of using sponge iron as a precipitant for the copper in solution resulting from desliming of the ore in a dilute sulphuric acid solution. Anaconda had done considerable work on the production of sponge iron in 1914 for use as a precipitant of copper from leach solutions. Some success and considerable experilence were attained at the time. indicating that, sponge iron might be successfully made by a modification of the process used in 1914, a batch process in which an iron calcine was reduced by means of soft coke, resulting from noncoking coal, in a Bruckner-type revolving horizontal cylindrical furnace widely used 50 years ago. The coke and calcide formed the bed in the Bruckner furnace, which was rotated at about 1 rpm. The bed was brought to a temperature of about 1800°F by means of an oil flame over the surface. Although results were reasonably satisfactory, they did not warrant full development of the process at that time. A good deal of work has been done in the last 50 years on the production of sponge iron. The objective in some cases has been the production of a precipitant for copper from solution, but the bulk of the work has been done for the production of open-hearth steel furnace stock. The production of an open-hearth stock presents two problems rather than one: first, producticon of the sponge iron, and second, what is perhaps of equal difficulty and importance, conversion of the sponge iron into a form suitable for use in the open-hearth furnace. So far as is known to the writer, none of the sponge iron processes tried in the past have proved to be economically feasible. However, Anaconda had a combination of conditions appearing to justify an attempt to produce sponge iron which would serve for the leach-precipitation-float process. It was thought that the process used in 1914, if changed to a continuous one, might work out satisfactorily. The following favorable conditions at Anaconda justified the investigation: 1—A sufficient tonnage of good grade iron calcine resulting from the roasting of a pyrite concentrate in one of the acid plants, at substantially no cost. 2—Reasonably cheap natural gas. 3-—The fact that there was no need for production of a high grade product. 4— The fact that there was no need for obtaining a consistently high reduction of' the iron in calcine. A small revolving Bruckner-type furnace about 2 ft ID by 4 ft long was set up for early pilot work at the research building. This pilot furnace showed that a satisfactory product could be obtained at reasonable cost. It also indicated a marked advantage in preceding the reduction furnace with a furnace of similar size and capacity for preheating and roasting out any residual sulphur from the feed. The small furnace was operated for several months, various details of the process were worked out. and sponge iron was produced to supply a pilot LPF plant which treated 300 lb of Kelley ore pel- hr. Later a second pilot furnace 5 ft in diam and 12 ft long inside was set up at our reverberatory furnace building. This furnace confirmed the data of the small furnace and gave a basis for design of the final plant. At Anaconda a pyrite concentrate, running about 48 pct S, is recovered from copper concentrator tailings by flotation. This concentrate is roasted to sulphur of 3 pct or less at the Chamber acid plant. The iron calcine contains about 57 pct Fe and 18 pct insoluble. The iron calcine feed, as mentioned before, is re-roasted and preheated in a reroast furnace preceding the reduction furnace. Both are of the Bruckner type. The reroasted calcine is fed into the reduction furnace at 800" to 1000°F along with 30 pct slack coal. In the feed end of the furnace the volatile is burned from the slack, giving a soft coke which readily serves for reduction of the iron. Hard metallurgical coke will not serve the purpose. since it does not reduce CO readily at a temperature of 1850°F. All indications are that the actual reduction of the iron is accomplished by carbon monoxide below the surface of the bed, which is 30 in. deep at its center. Apparently there is a constant interchange: Fe²O³ + 3CO = 2Fe - 3CO², CO² + C = 2CO Actually iron oxide is reduced by CO at somewhat lower temperature than the 1850 °F used in the process. but this temperature is necessary to obtain a satisfactory rate of furnace production. The furnace atmosphere is generally reducing, and typical blue carbon monoxide flames satisfactorily cover the bed. Gas flames from four 3-in. Denver Fire Clay Inspirator burners are played directly on the bed, which is slowly cascaded by the 1 rpm of the furnace. An excess of coke is necessary to assure maintenance of good reducing conditions in the furnace bed. Part of this coke is recovered for re-use.

Jan 1, 1954

By Herbert Cremer

The United States,-the largest nickel consumer in the world, produces virtually no nickel from domestic ores. Dependence upon outside nickel - sources and increasing strategic, importance of nickel focused the attention of the United States on deposits within its boundries. The most likely domestic sources of nickel are low-grade oxide and silicate deposits, such as the Cle Clum deposit, Kittitas County, Wash., and the Riddle deposit, Douglas County, Oreg. The Federal Bureau of Mines, Northwest Electrodevelopment Laboratory, Albany, Oreg. has been investigating electric smelting of low-grade oxide and silicate nickel ores. These smelting studies have been on selective reduction to produce low-carbon ferronickel. Selective reduction is based upon the relative reducibility of: nickel, iron, and chromium, in the order named, when the charge contains insufficient carbon or other reductants to reduce the entire quantity of these metals present in the ore.

Jan 1, 1954

By John D. Price

R. W. Campbell (Jones and Laughlin Steel Carp., Pittsburgh)—As usual John Price has presented an excellent paper. I know of no one who has devoted more time and conscientious thought to this subject than he has, and his efforts have produced scientific data and conclusions of real value. Results of similar studies made by the Jones and Laughlin Steel Corp. on the blending of Pennsylvania coals may be of interest. During 1935 the late F. W. Wagner and the present writer developed an empirical physical fuel value formula for evaluating the combined physical tests of metallurgical coke, see Table IV. In the development of the formula, the tumbler test, shatter test, screen test, and porosity of coke have been taken into consideration. Each of these tests brings out some physical property not expressed by the others, and in estimation of the importance of these tests as a medium for evaluating coke for the blast furnace, their worth has been graded as follows: It was thought that some additional penalty should be applied to cokes that do not show a uniform screen size, assuming that coke larger than 4 in. was not desirable and that this size was an indication of irregular size. Therefore the screen test factor is further corrected by applying a penalty of 1 pct for each 1 pct material greater than 4-in. size as shown by the screen test. The physical fitness of coke for blast furnace use depends upon four fundamental characteristics: l—gen-era1 strength, as indicated by the shatter test; 2— abradability, or the resistance of cell walls to reduction in size by attrition, as indicated by the tumbler test; 3—porosity; and 4—size and uniformity. Each of these characteristics is not of equal importance, but surely, in an attempt to determine an accurate physical value, the individual value of each should be taken into consideration. The formula is used as a yardstick, so to speak, in determining variations in the normal production of metallurgical coke and in making comparisons between normal coke and experimental blends with other coals. Table IV gives results of coking tests on raw and washed high volatile coals and these same coals with blends of coke breeze, anthracite coal, char and low volatile coals, all from Pennsylvania. The table also gives the source of the coals and coke breeze, as well as their proximate analyses. It will be noted that the washing of the raw high volatile coal increased the physical fuel value from 26.17 to 39.16, showing the effect of removing shale and slate which produces a detrimental fracturing and weakening of the coke. It is a well known fact that many Pennsylvania high volatile coals contain an excess of resinous material or binder which has a tendency to weaken the coke, owing to its high evolution of gases during carbonization. Tests were therefore made with coke breeze, an inert material, in an attempt to alleviate this condition. Blends of washed coal with 5 pct and 7.5 pct coke breeze collected from the coke plant quenching sump (82.9 pct through 1/8-in. screen) resulted in cokes with a physical fuel value of 20.75 and 16.73 respectively, lower than results obtained from 100 pct raw high volatile coal. It will be noted that the percentage of + 4-in. coke increased but that the tumbler test was very low. Undoubtedly better results could have been obtained if the coke breeze could have been crushed extremely fine. During World War II the government asked the coking industry to consider the use of anthracite coal as a substitute for high quality low volatile coal, which was extremely scarce at the time. The next series of tests was made with blends of raw and washed high volatile coals and 5, 10, and 15 pct Pennsylvania No. 5 buckwheat anthracite (100 pct through l/8-in. screen). It was necessary to use some raw high volatile coal in these blends because of insufficient washer capacity at the time. It will be noted that the physical fuel values

Jan 1, 1954

By D. F. Gibbons

THIS investigation was undertaken in order to gain further information on the mechanism of plastic deformation of iron at temperatures between room temperature and 77.2°K, and also to contribute to our understanding of the low temperature brittleness phenomenon that occurs in both the tensile and impact tests. It was decided to study as pure an iron as possible and to use the tensile test in order to simplify the problem as much as possible. Two series of iron were used throughout the investigation: iron A, obtained from the National Research Corp. and supplied as gas-free, high purity iron; and iron B, obtained from Fast' and produced by melting iron under purified hydrogen. Typical analyses of samples of both irons, after fabrication, are given in Table I. The main difference in composition of these irons is in their nickel and oxygen contents. Nickel has a very large solid solubility in iron. However, oxygen has an extremely small solid solubility. An attempt was made to reduce the oxygen content still further by remelting under purified hydrogen. This was successful in reducing the oxygen content to -0.0006 pct. However, no results on tests of this iron are included, since difficulty was encountered in fabricating the ingot. Wire specimens 1 mm in diameter were used for the investigation. The wires were produced from the ingots, as supplied, by cold rolling and swaging, with intermediate vacuum anneals at 700°C. Care was taken to reduce preferred orientation in the wires to a minimum by keeping the percentage of reduction in area between anneals of 30 to 40 pct. X-ray transmission photographs showed no indication of preferred orientation. The final grain size of iron A was 20 to 24 grains per linear mm and of iron B, 16 to 18 grains per linear mm. All specimens were tested in the decarburized and denitrided condition unless otherwise stated. The procedure was to pass hydrogen, saturated with water vapor at 26oC, over the specimens at 730°C for 2 hr. The specimens were quenched in iced water and then annealed in a vacuum of - 5x10-6 mm Hg for one day at 560°C to remove hydrogen. Higher temperatures were used for the vacuum anneal but this temperature was found to be quite satisfactory in removing hydrogen. The specimens were slow-cooled after the vacuum anneal. Where it is stated that the specimens were carburized, the specimens were given an identical decarburizing treatment and then carburized by passing hydrogen, saturated with pure n-heptane vapor at 0°C, over the specimens at 730°C. The specimens were then homogenized at 730°C for 2 hr in dry hydrogen. After the carburizing treatment the specimens were given the same vacuum anneal as for the decarburized specimens. The carbon content was determined from the maximum value of the internal friction carbon peak.2 Apparatus Fig. 1 shows a diagram of the tensile apparatus, with half the cooling coil removed. The cooling coil enabled temperatures from 200" to 77.2oK to be obtained. The apparatus was designed to fit a standard Tinius Olsen Universal tensile machine. The apparatus consisted of the rigid plate A which carried two carefully machined hooks H. These hooks attached the plate firmly to the moving crosshead of the tensile machine. The lower grip B was rigidly suspended from the plate A by three stainless steel rods. The upper grip C was joined to the upper grip of the tensile machine by the stainless steel rod E. The lower part of the apparatus was surrounded by a Dewar vessel, which was made a sliding fit inside the brass collar D. The specimen was held in a clamp which fitted into the grip, forming a ball and socket joint (see inset a). This was to attain as nearly pure

Jan 1, 1954

By Harold J. Rose

It is a real pleasure to attend another meeting of the Institute. This is partly because I am a Westerner by birth and training. As a young man I had personal experience with firing Colorado and Wyoming coals. The coal industry of the United States has done an outstanding job of producing coal. We can all be proud of the industry's record of continuously adopting technical improvements in coal mining. NEED FOR MORE EFFECTIVE COAL MARKETING However, it cannot be said that the industry has done an equally good job of marketing coal. There has been a rapid growth in U. S. population, and an even greater expansion of industrial demands for heat and power. Yet coal consumption for the past several years has been no greater than it was 40 years ago, in 1912 and 1913.

Jan 1, 1954