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By Lawrence Henderson

TO increase tonnage in the early days of coal mining it was necessary only to hire more men. The job now is to increase the tons per man, but other troubles arise because this has been accomplished. With the introduction of continuous miners, newer and faster loading machines, and faster belt conveyors the safety engineer has been faced with new difficulties. Use of larger equipment created roof control problems, which were thoroughly studied by the safety engineers of various companies. With the assistance of the U. S. Bureau of Mines, safety engineers in Alabama conducted roof bolting experiments wherever they appeared to be applicable. Accident records before and after the method was adopted showed results more than justifying the efforts and funds expended. The larger mining companies responded enthusiastically to the program because roof bolting not only prevented accidents but also increased production efficiency, in many cases rep- resenting the margin between profit and loss. In Alabama today there are miles of roof supported by bolts.

Jan 8, 1957

By J. M. GARZA

The production of oil in Mexico during 1945 was 43,402,852 bbl., or a daily average of 118,912 bbl. In March 1938 most of the oil properties were taken over by the Mexican Government and since then have been operated by Petroleos Mexicanos, an agency of the Government. The variations in the production during the last 8 years have been due to- fluctuations in the export market demand for Mexican oil rather than to the productive capacity of the fields. Since 1937, yearly production, to the nearest hundred thousand barrels, has been as follows :

Jan 1, 1946

By H. H. Klepfer, M. E. Snyder

UNTIL very recently, the commercial availability of the rare earths as metals has been very limited. Fabrication of mill products from these metals has not been studied in most cases. This note reports the results of the development of fabrication techniques for thulium. Thulium has a melting point of about 1550°C and a hexagonal close-packed crystal structure. It oxidizes in air to give a black oxide (TmzO,). The procedures for producing thin thulium foil were developed on one ingot weighing about 220 g and were subsequently applied in processing about a pound of metal to foil. Several alternates for the various fabrication steps were investigated and will be discussed. The metal fabricated was in the form of commercial chill-cast ingots 1 in. in diam and 2 in. long weighing approximately 220 g. Impurities in the ingots were reported by the vendor to be 4000 ppni tantalum, 2000 ppm calcium, 200 ppm nickel, 100 to 200 ppm iron, 100 ppm europium, and less than 100 ppm copper, lutetium, and ytterbium. In addition to these impurities, several salt-like inclusions as large as l/8 in. in diam were revealed along the center line of the one ingot sectioned. Preliminary tests indicated that small wafers cut from the as-cast ingot would not fabricate readily by rolling. Forging of copper-jacketed wafers was therefore attempted. At 1550"F forging was satisfactory but an apparent reaction of copper with thulium demanded investigation of lower temperatures. Therefore, the remainder of the test ingot was forged at 1450°F—with only minor cracking. All ingots forged were inserted into copper tubes of 1 in. 1D and 0.125-in. wall thickness. The jackets were sealed by flattening the ends of the tubes and welding under helium. Heating time at 1450°F was 30 min. Press forgings of 1/8 in. per pass were used, followed by 10 min reheats. When the ingots had been squared and reduced to 0.250 in. in thickness, the original copper jacket was stripped off and replaced by a new jacket in preparation for hot rolling. Hot rolling at 1450" F without edge cracking was readily accomplished after forging. Excellent results were obtained with 10 pct reductions of thickness followed by 5 to 10 min reheats. After reduction of thickness from 0.250 to 0.100 in. the copper jacket was removed. It was found, in fact, that hot rolling in air was possible. A tenacious black oxide similar to that seen on zirconium was formed during 3 min reheats at 1450°F. Reduction in air to 0.010 in. foil was possible taking 10 pct reductions per pass. The oxide coat formed during hot rolling in air could best be removed by sand blasting and pickling. Common pickling solutions containing polar solvents were found to attack the metal too rapidly and a concentrated nitric-hydrofluoric acid mixture attacked neither the oxide nor the metal. The most satisfactory pickling solution was 52 vol pct concentrated nitric acid-48 vol pct glacial acetic acid. After forging to 0.250 in., vacuum annealing and cold rolling was found to be another satisfactory alternate to hot rolling in copper jackets. After forging. a hardness of Rockwell B76 was found. Annealing in vacuum (2 X l0-5 mm of Hg) for 1 hr at 1200"F did not alter this value. Annealing at 1470"F for 1 hr brought the hardness down to R;]63. With cold rolling (5 pct per pass) the hardness returned to about R,1:76 after 20 pct reduction and edge cracking became noticeable. However, cold rolling to a total of 40 pct reduction in thickness (RI,,83) was possible before edge crack propagation became serious. Good surface finish was obtained, and the metal loss due to oxidation was minimized by cold rolling and vacuum annealing. Using this procedure the yield of 1.25 in. wide by 0.010 in. thick foil from a 1-in. diam ingot was about 40 pct.

Jan 1, 1961

By L. W. Holm, G. G. Bernard

Laboratory experiments were conducted to determine the effect of foam on gas flow in porous media. Previous studies have indicated that foam may be applicable as a restrictive agent in influencing underground gas flow. Foam was found to be exceedingly effective in reducing the permeability of porous media to gas. Consolidated and unconsolidated sands with specific permeabilities of 100 to 146,000 md had, in the presence of foam, gas permeabilities that were less than 1 per cent of the specific permeability; in many cases the gas permeability was practically zero. Foam reduced the gas permeability of loose sand to a much greater degree than that of a tight sand. For example, the permeability of a 125,000-md sand was reduced to 3 md while the permeability of a 4,000-md sand was reduced to 7 md. This effect should cause, to some degree, selective plugging of high permeability channels in various oil displacement processes. The presence of oil in a porous medium decreased the effectiveness of foam in reducing gas permeability; apparently oil acts as a foam depressant. However, it was found that certain foaming agents were very effective in reducing permeability even. in the presence of oil. Also, continuous injection of other foaming agents increased their effectiveness, when oil was present. The effect of foam on permeability of porous media to gas was studied as a function of foaming agent concentration and injection rate, absolute permeability, total pressure, pressure gadient, length of porous system, brine concentration and time. INTRODUCTION The use of surface active agents in flood water to increase the recovery of oil has been studied in the laboratory and in the field for decades, with rather limited success. In recent years a new approach to the problem was proposed. Instead of using an aqueous solution of surfactant as an oil recovery agent, Bond and Holbrook2 proposed that the oil recovery agent be a mixture of surfactant solution and gas. In their method, a water-soluble surface active agent with foam-producing characteristics is injected into an underground formation as an aqueous slug. This slug is followed by gas to produce a foam within the rock. Foams are defined as "agglomerations of gas bubbles separated from each other by thin liquid films'! A foam is fundamentally an unstable system. The foam process for oil recovery has since been studied by other investigators. Fried 1 has shown that foam can displace from porous structures oil that normally is unrecoverable by conventional water or gas drives. This superior oil displacing action is believed to be the result of several factors: (1) foam introduces into the reservoir many resilient interfaces of various sizes and curvatures, which increase the probability that a proper combination of forces for oil displacement will be created; (2) foam has appreciable viscosity which improves mobility ratio and contact efficiency; (3) foam accentuates the trapped gas effect because high gas saturations are possible without producing high gas/oil ratios. Deming4 studied the effect of various foam properties on the displacement of liquid. He found that (1) high foaming ability favors high displacement efficiency; (2) high foam stability is not necessary for high displacement efficiency; (3) displacement efficiency decreases with increase in plasticity of foam and (4) displacement efficiency is unaffected by surface tension of foaming agent solution. One of the important aspects of this oil recovery process is the effect of foam on gas permeability. The simultaneous flow of gas and liquid in porous media has been studied by numerous investigators and a large amount of literature exists on the subject. In this previous work, gas and liquid have generally been considered as essentially independent phases, whose flow characteristics are related through the saturation parameter. Foam, however, is a material with properties that are considerably different from those of its components; for example, the viscosity of a foam is greater than either of its components. Gas flow in a porous medium in the presence of foam should be considerably different

Jan 1, 1965

By J. C. Holm

ALTHOUGH the main constituents of Portland cement are the oxides of calcium, silicon, aluminum, and iron, characteristics of the cement are seriously affected by such contaminants in the raw materials as magnesium, phosphorus, sodium, potassium, sulfur, and others of lesser importance. Sources of cement raw materials comprise a long list of minerals occurring in nature in various degrees of purity. The calcareous raw materials are generally limestone of purity varying from calcite to cement rock (a low-grade limestone of natural, almost correct chemical composition), chalk, marl, seashells, coral, and coral sand. The argillaceous raw materials supplying silicon, aluminum, and iron are mainly various forms of clay and shale, to which may be added silicious sand, iron ore, bauxite, or kaolin for minor adjustment of the chemical composition. The quality of Portland cement in this country, as in most parts of the world, is controlled by rigid specifications limiting both its chemical and physical characteristics, and cement manufacturers must comply with these specifications to market their products. Brandon Raw Materials: An unusual deposit of marl, limestone, and clay occurs 12 miles east of Jackson, Miss., near the village of Brandon. The deposit is of sedimentary origin and of relatively young geological age. Layers of marl vary from putty-like consistency to such hardness that it can be cut with a knife. Interspersed between the layers of marl are bands of limestone varying from very soft to hard. the compressive strength of the hardest layers exceeding 15,000 lb per sq in. The layers of marl range in thickness from a few inches to about 25 ft. The limestone bands seldom exceed 2 to 3 ft but often occur in thinner veins in the quarry face. Above the top layer of the deposit, which may be marl or limestone, a layer of clay ranges in thickness from a few inches to more than 20 ft. Part of this clay is harvested with the raw material to supply part of the argillaceous components, but most of it is stripped as overburden and wasted. A highly silicious clay of extreme fineness, harvested from the bottom of the quarry below the marl and limestone bands, is to correct the silica content. Fig. 1 shows typical examples of the interrelation of the various materials, based on drill logs obtained during exploration of the deposit. The marls vary in moisture content in the bank from 15 to 30 pct and contain up to 70 pct -200 mesh particles and up to 40 pct —10 p particles. A large percentage of the fine fraction is within the colloidal range, making the marl extremely sticky. The limestone beds range in hardness from that of the hardest marl to a medium hard limestone and in moisture content from 2 pct to approximately 14 pct. The line of demarcation is often difficult to draw and the moisture content is possibly the best means of classification. The ratio of soft to hard material is not constant throughout the deposit, but the average is approximately 70 pct soft to 30 pct hard material. Through the years this deposit had been repeatedly explored by cement companies and rejected as unsuitable for cement manufacture, mainly because of physical characteristics. In 1950, after careful exploration of the deposit and comprehensive laboratory and pilot plant tests of the raw materials, officials of Marquette Cement Mfg. Co. of Chicago were convinced that a practical and economical method of processing the raw materials had been devised and authorized construction of a new plant that began production in December 1951. Marl and limestone, common raw materials for Portland cement, usually occur in such a manner that they can be worked separately. Their harvesting and processing normally have presented no difficult problems, as standard equipment and methods are available for handling either sticky or hard materials. No other cement plant utilizes a deposit of the nature of that at Brandon, however, and for this reason no prior art was available for design. When properly blended, the raw materials produced an excellent cement without additives. Basic problems to be solved in designing the cement plant were:

Jan 1, 1957

By F. W. Archibald

The American Smelting and Refining Co. has standardized its copper converting practice to attain a maximum unit blister production with a minimum of refractory consumption by careful location of the tuyeres and by applying magnetite coatings on the hard-burned magnesite brick linings. THE American Smelting and Refining Co. operates four primary copper smelters in the United States with a total of 17 Peirce-Smith type converters; 15 of them are 13 ft in diameter by 30 ft long, and two are smaller. Some details of operations vary with locale; however, fundamentals of design, operation, and maintenance are common to all plants. All converter shells are l-in. thick except for one new converter with riding rings on the ends which has a shell thickness of 1 1/2 in. More tuyeres can be installed with rings on the ends and hence more air can be used. Welded construction is replacing riveted. Minor shell repairs are made after each campaign. Principal causes of complete shell replacement are warpage and cracking resulting from localized over-heating. A 13x30 ft shell is being replaced after 34 years of operation and a total production of approximately 750,000 tons of blister. Converters are driven by 80 hp DC motors through company-designed worm gear reducers. Rotation is 0.38 rpm which permits the skimmer to spot the converter quite accurately for skimming slag. At three of the plants, protection against unscheduled air or power failures is afforded by the installation of emergency drives, consisting of auxiliary air motors or storage batteries. In order to avoid excessive splash out of the converters, the converter mouths are located as far back on the shells as existing flue facilities will permit. At one plant, the back of the mouth is only 13" to the rear of the vertical center line of the converter,. whereas it is 28" at another plant. Newly-lined mouth areas vary from 36 to 44 sq ft with effective operating areas about 25 pct less. It is important to keep the converter mouths as clean as possible. Dirty mouths create back pressures in the converters and as a consequence the tuyere air volumes are reduced. Mouths are generally cleaned by bumping with an empty ladle. Small mouth-cleaning rams have been used but unless extreme care is exercised the brickwork may be damaged. After considerable experimentation, the plants standardized tuyere elevations at 4 to 4% in. below horizontal center line at the shell with a downward pitch of about 13/16 in. per ft of length. Tuyeres are spaced at 6 in. centers except at the riding rings where there are no tuyeres. With this tuyere location, several of the plants now freeze magnetite slag in the bottoms, fronts, and backs up to the bottom of the tuyeres to control the internal shape of the converter. This has the effect of improving the agitation and mixing so that there is a marked increase in converting speed besides affording protection to the brick lining. Currently, the trend is to increase tuyere diameters from 1% to 2 in. to increase the air flow. At present, all converters are hand-punched but one converter is being equipped with a set of mechanical tuyere punchers. Rods for punching are % in. hexagonal smelter bar upset to ll/s in. and rods for cleaning tuyeres are upset to about 1% in., or larger, depending upon the tuyere diameter. Two of the plants use pneumatic reamers for cleaning the tuyeres between charges to minimize disturbing the coating on the inside of the refractory lining. Most of the puncher's platforms are pneumatically or hydraulically mounted so that a convenient punching position can be maintained regardless of tuyere position. For normal operations, tuyere air pressures vary from 15 to 13.5 psi, although some cycles such as magnetiting require pressures down to 8 psi. Air requirements vary from an average of 25,000 cfm on the newer installations down to 12,000 cfm on the older ones. Converter hoods are designed to protect the punchers from sparks, splash or hood accretions as well as to prevent the escape of objectionable

Jan 1, 1955

T.B.King (Depaytment of Metallurgy, Massachusetts Institute of Technology)— A valuable contribution of the authors is in the factual information which they have been able to gather; this type of information is quite difficult to obtain. In many respects, however, it would have been better if they had not subsequently embarked on a discussion of the chemistry of the converter process. It seems inconceivable that the authors do not refer to the papers of Schuhmann and his associates14 which have set the thermodynamic foundation for the whole copper smelting operation. In addition, a very useful review on the physical chemistry of copper smelting by Ruddle 5 appeared as long ago as 1953. An examination of this literature would have convinced the authors that there is no cause to be surprised at a correlation between the magnetic content and the silica content of converter slags, though they rightly point out that one should distinguish between the total magnetite content of the slag and the amount of magnetite which may be considered to be in solution. It is not true that the lowest melting converter slag is that corresponding to the eutectic between ferrous oxide and silica. The simplest slag system which can be considered is a three-component system, since both ferric and ferrous iron are present. As Schuhmann, Powell, and Michal have shown, there are lower melting compositions than this eutectic in the ternary system. The most unfortunate impression given by this paper is that the driving force for chemical reaction is determined by the heat of reaction: of course the entropy change must be taken into account. It would have been more correct to list, in Table VI, values for free energies of formation. Nor can it be said that the data in Table VI represent the "best available data." They do not corregtond with any of the recent, acknowledged sources. F. E. Lathe and L. Hodnett(author's reply)— We are pleased that Dr. King finds the factual information in our uawer of some interest. Dr. King suggests that it would have been better if our analysis and discussion of the data had been omitted, largely because our list of references is so. incomplete. If he will carefully read our introduction, he will see that the questionnaire was sent out in the hope of obtaining data which would throw light on certain questions relating to the use of converter refractories. We did not attempt (nor would the AIME have published!) a complete review of the literature on copper converting, as Dr. King has apparently assumed, nor indeed a complete analysis of the data submitted, but tried only to find a sound basis for the choice of refractories, taking into consideration common variations in converter practice. We hope our paper indicates that, by raising the silica content of the converter slag and operating at a higher temperature, the normal circulating load of magnetite can be greatly reduced, and the whole reverberatory-converter operation improved to a major degree, with resultant important savings. Under such operating conditions, chrome-magnesite brick may be expected to stand up better than those of straight magnesite. Regardless of the choice as between these brick types, however, we find the cost of converter refractories to be so low in comparison with other converter costs as to justify operation under the more severe conditions suggested. Valuable as are the papers by Schuhmann and associates and the book by Ruddle, we make no apology for omitting reference to them, nor for using heats of reaction without mention of entropy changes or free energies of formation. Our primary object was to interest the practicing copper metallurgist, with whose language we may claim to be fairly familiar; we think it would have been unwise to include the highly theoretical phases of the subject which Dr. King suggests. The interest shown in our preliminary paper presented at the New York meeting in 1956, and the trends in practice whtat we have observed since that time, suggest that we did not wholly miss the target. In conclusion, we sincerely hope that Professors King and Schuhmann will independently review the data obtained in our questionnaire and submit a paper giving their own recommendations as to the choice of refractories and the particular converter operating conditions which will result in the lowest overall cost of copper smelting and converting.

Jan 1, 1960

By G. L. Simard, D. J. Salley, J. Chupak

DITHIOPHOSPHATES and xanthates are the principal collectors for sulphide minerals, and consequently any knowledge of mineral-collector systems of this type is of value. In the present investigation an attempt was made to obtain information on the interaction of a typical dithiophosphate with galena. In carrying out this study, radioiso-topes, so much discussed in the past few years,', ' were extensively employed. By using radioactive dithiophosphate synthesized from radioactive phosphorus, a rapid and sensitive analytical procedure G. L. SIMARD is in the Research Division, Stamford Research Labs., American Cyanamid Co., Stamford, Conn.; J. CHUPAK, formerly in the Research Division, Stamford Research Labs., is now at Camp Detrick, Frederick, Md.; and D. J. SALLEY is in the Research Division, Stamford Research Labs. AlME New York Meeting, Feb. 1950. TP 2815 B. Discussion (2 copies) may be sent to Transactions AlME before April 30, 1950. Manuscript received Oct. 17, 1949. was at hand. This permitted determination of such important quantities as the rate of uptake of dithiophosphate by the mineral, the amount existing at equilibrium on the mineral surface and in the solution, and the desorption of the agent from the surface. Such measurements were possible even on single crystals of galena. In addition the exchange of dithiophosphate between the solution and the sorbed phase was examined, a matter which could be accomplished only by the use of isotopes. It is desired to point out at this time that throughout this paper the term "sorption" has been employed to designate the uptake of agent by mineral, without implication as to the nature of the process by which the uptake was accomplished. Experimental Materials: Galena: The galena was from the Tri-State district. For studies on ground mineral four preparations were used during the course of the investigation. These were prepared from selected large crystals by wet grinding in order to reduce surface oxidation. Alcohol was chosen as a convenient medium for this purpose. The ground mineral was then fractionated by sedimentation in alcohol, dried by evacuation, and stored in a nitrogen-filled desiccator. Data on the preparations are tabulated in table I. The size analysis indicated that the areas of the preparations were of comparable magnitude, even though the absolute values may be somewhat incorrect. Dithiophosphates: Nonradioactive dithiophbsphate (di-isopropyl or di-secondary butyl) was obtained by purification of a commercial product. An aqueous acid solution of the agent was extracted with petroleum ether, the ether layer dried, and the dithiophosphate precipitated as ammonium salt with anhydrous ammonia. Several repetitions of this process resulted in a nearly colorless, flaky product of good purity. (General formula (RO,)PSSNH,). Radioactive dithiophosphate was synthesized by heating radioactive elementary red phosphorus* • For early experiments (1943 to 1944), the elementary red phosphorus was obtained from the cyclotron group at the Crocker Radiation Laboratory of the University of California. Berkeley, through the courtesy of Dr. Joseph W. Hamilton. For more recent work (1946 to 1949). the elementary radioactive Phosphorus has been supplied by the Oak Ridge National Labarotories on allocation from the U. S. Atomic Energy Commission. with sulphur at 270" to 300°C to produce radioactive P,S,. The latter was then treated with the desired purified alcohol (isopropyl or secondary butyl) at 60" to 80°C to form the dithiophosphoric acid derivative. Purification was effected in the same manner as for the nonradioactive material. The effectiveness of the purification method was established by the isotopic dilution method." The same technique was used to show that only negligible decomposition of neutral or of carbonate solutions of the dithiophosphate took place over a period of a day; this was true whether or not galena was suspended in the solutions. Procedures: Radioactivity: The activity of a solution was obtained by counting with a small glass-jacketed, silvered Geiger counter, using a conventional scaling circuit.' Crystals and other solids were counted under a bell-shaped, mica-window counter tube. Sufficient counts were made so that the probable error of counting was of the order of +I-3 pct. The specific activity in terms of counts per minute

Jan 1, 1951

By R. T. Hukki, P. Voutilainen

This paper describes a new apparatus for continuous wet fine-screening. Its preferred size range seems to be from 0.1 to 1 mm. The supercritical trommel screen is a short cylindrical trommel of wedge-bar type construction operated at a speed greater than the calculated theoretical critical speed and at a speed lower than the speed at which particles introduced into it will centrifuge. The paper describes results obtained with the prototype supercritical trommel screen in pilot plant scale. The results point to a high capacity and simultaneously to a high sharpness of screening. Finally, an analysis is presented of the forces acting in screening on a supercritical trommel screen. In the era of high speed, high capacity processing of raw materials the conventional slowly rotating trommel screen is often considered a relic from the times passed. A typical trommel screen is the cylindrical trommel rotated about a horizontal or slanting axis at a relatively low speed. Its screening characteristics are well known. Trommel screens of various types of construction, methods of operation, performances and other characteristics have been described in detail by Taggart.1 The theoretical critical speed of a trommel screen is defined by the same well known equation as that applied for grinding mills: where ncritical = the numerical value of the theoretical critical speed in rpm, D 1 = the inside diameter of the trommel in feet, and D2 = the inside diameter of the trommel in meters. ncritical corresponds also to the centrifuging speed provided that the coefficient of friction f between the trommel surface and the outer layer of the charge is = 1.0. In a general case: From Eq. 2 it is seen that in all cases where f < 1.0 the centrifuging speed of the trommel is greater than the theoretical critical speed. The supercritical trommel screen concerns screening operations within that speed range which is between the theoretical critical speed defined by Eq. 1 and the centrifuging speed defined by Eq. 2, under conditions such that the coefficient of friction f<1.0. It is reported by Taggartl that the optimum speed of a conventional cylindrical trommel screen 36 in. in diam was experimentally found to be about 16 rpm. From Eq. 1, the critical speed of this trommel is 44.2 tpm. The indicated optimum speed corresponds to (16/44.2) x 100% = 36.2% of the theoretical critical speed. The optimum speed range based on information available seems to vary from 33% up to 45% of the theoretical critical value. The trommel screen operated at a supercritical speed: operates at a speed greater than the calculated theoretical critical speed; operates at a speed lower than the speed at which the charge introduced into it will centrifuge; operates within the supercritical speed range in such a way that the coarse fraction of the charge must slide in respect to the inside surface of the trommel; the outer layer of the said fraction next to the screen tends to proceed in the same direction as any selected point on the trommel surface, but now at a lower speed, and indeed at a subcritical speed; allows to a certain degree adjustment of the maximum determining particle size x of the fine product simply by a variation of the speed of the trommel within the supercritical speed range. With increasing trommel speed, finer and finer material only will get the opportunity to find its way into the fine product while the screen opening remains constant. In conventional practice, the low speed trommel is generally used for sizing of coarse and medium coarse material, but seldom for fine screening. The characteristic feature of the product passed through the conventional trommel is that the determining size x of the biggest particle is barely smaller than the nominal size of screen opening. The supercritical trommel screen, on the other hand, is best suited for sizing of relatively fine material. Furthermore, the characteristic determining size x of the biggest particle in the fine product can be substantially smaller than the nominal width of the screen opening or slot. In the trommel screen best suited for the applications

Jan 1, 1965

By R. J. Wagner, G. H. Gronhovd, A. J. Wittmier

The paper describes the results of a series of full-scale boiler tests run to determine the ash fouling characteristics of a North Dakota lignite as a function of sodium content of the coal. Four levels of sodium were investigated and rate of fouling was determined both by boiler performance and by specially constructed probe equipment. Discussed is information on chemical and physical properties of the ash deposits. The vast lignite deposits in North Dakota are becoming increasingly important as a source of low-cost fuel for power generation. A new 200-megawatt lignite-burning plant located near Stanton, N.D., was placed in operation early in 1966. Another 172-mega-watt unit in the same area is scheduled for operation late in 1966, and plans have recently been announced for a 212-megawatt unit to be in operation by 1970. If these plants show good performance, western North Dakota could become a major power production center for a large geographic area in the upper midwest. One of the problems associated with the use of lignite in power boilers is its tendency to form troublesome ash deposits on boiler tube surfaces. This problem varies considerably from one lignite-fired boiler to another and is dependent upon such factors as source of lignite, boiler design, load factor, tube metal temperatures, and soot-blowing practices. This report presents data on sodium content* variation in a North Dakota mine, describes the results of a series of tests performed to determine the relationship between ash fouling tendencies and the sodium content, and tells how the mining company and electric utility have coordinated their operations to reduce the ash fouling problem. GENERAL BACKGROUND Until recent years it was generally assumed that North Dakota lignite from a given mine had qualities peculiar to that mine, but that within the mine, the lignite was quite uniform. In 1959, the Otter Tail Power Co. installed a 53.5-megawatt lignite-burning unit at its Hoot Lake plant, Fergus Falls, Minn. Shortly after the unit was put into operation, it was noted that at certain times under high load conditions, ash fouling problems were quite severe, but not consistent or predictable. Each time fouling occurred in the unit the position of the loading shovel at the mine was noted. Within two years it became apparent that there were certain areas in the Knife River Coal Mining Co. mine at Beulah, N.D., where the lignite apparently had a high fouling tendency for the Hoot Lake unit. Results from analyses of the coal did not show large differences in moisture percentage, ash percentage, heating value, or ash fusion temperature between the highly troublesome and less troublesome coals. After three years of operation, the troublesome areas had been quite well defined and the mining company started a procedure of bypassing these areas when load conditions at the plant were high and then returning to the troublesome areas to load the lignite during the summer months when peak loads were lower. It was noted that the lignite from the areas which caused severe fouling under high load conditions produced much less fouling at the lower summer loads. However, bypassing certain areas in the strip pit caused the mining company to disrupt their continuous loading schedule, which in turn upset their continuous stripping schedule and made for less efficient mining operation. Meanwhile, at the Hoot Lake plant, the number of soot blowers on No. 2 boiler had been increased from 30 to 55, and a very rigorous blowing schedule had been adopted resulting in soot blowers operating at least some part of 21 hrs per day. Fouling problems continued, with a resulting reduction in plant economy and reliability. During this period the Federal Bureau of Mines Coal Research Laboratory at Grand Forks, N.D., had initiated various research programs relating to the study of lignite ash and ash fouling problems. One project was a survey in which full seam samples were obtained from various locations in all the major

Jan 1, 1968

By A. Stanley, J. C. Mead

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

Jan 1, 1950

By Joseph C. Mead, Alan Stanley

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

Jan 1, 1950

By J. Chupak, D. J. Salley, G. L. Simard

DITHIOPHOSPHATES and xanthates are the principal collectors for sulphide minerals, and consequently any knowledge of mineral-collector systems of this type is of value. In the present investigation an attempt was made to obtain information on the interaction of a typical dithiophosphate with galena. In carrying out this study, radioiso-topes, so much discussed in the past few years,&apos;, &apos; were extensively employed. By using radioactive dithiophosphate synthesized from radioactive phosphorus, a rapid and sensitive analytical procedure G. L. SIMARD is in the Research Division, Stamford Research Labs., American Cyanamid Co., Stamford, Conn.; J. CHUPAK, formerly in the Research Division, Stamford Research Labs., is now at Camp Detrick, Frederick, Md.; and D. J. SALLEY is in the Research Division, Stamford Research Labs. AlME New York Meeting, Feb. 1950. TP 2815 B. Discussion (2 copies) may be sent to Transactions AlME before April 30, 1950. Manuscript received Oct. 17, 1949. was at hand. This permitted determination of such important quantities as the rate of uptake of dithiophosphate by the mineral, the amount existing at equilibrium on the mineral surface and in the solution, and the desorption of the agent from the surface. Such measurements were possible even on single crystals of galena. In addition the exchange of dithiophosphate between the solution and the sorbed phase was examined, a matter which could be accomplished only by the use of isotopes. It is desired to point out at this time that throughout this paper the term "sorption" has been employed to designate the uptake of agent by mineral, without implication as to the nature of the process by which the uptake was accomplished. Experimental Materials: Galena: The galena was from the Tri-State district. For studies on ground mineral four preparations were used during the course of the investigation. These were prepared from selected large crystals by wet grinding in order to reduce surface oxidation. Alcohol was chosen as a convenient medium for this purpose. The ground mineral was then fractionated by sedimentation in alcohol, dried by evacuation, and stored in a nitrogen-filled desiccator. Data on the preparations are tabulated in table I. The size analysis indicated that the areas of the preparations were of comparable magnitude, even though the absolute values may be somewhat incorrect. Dithiophosphates: Nonradioactive dithiophbsphate (di-isopropyl or di-secondary butyl) was obtained by purification of a commercial product. An aqueous acid solution of the agent was extracted with petroleum ether, the ether layer dried, and the dithiophosphate precipitated as ammonium salt with anhydrous ammonia. Several repetitions of this process resulted in a nearly colorless, flaky product of good purity. (General formula (RO,)PSSNH,). Radioactive dithiophosphate was synthesized by heating radioactive elementary red phosphorus* • For early experiments (1943 to 1944), the elementary red phosphorus was obtained from the cyclotron group at the Crocker Radiation Laboratory of the University of California. Berkeley, through the courtesy of Dr. Joseph W. Hamilton. For more recent work (1946 to 1949). the elementary radioactive Phosphorus has been supplied by the Oak Ridge National Labarotories on allocation from the U. S. Atomic Energy Commission. with sulphur at 270" to 300°C to produce radioactive P,S,. The latter was then treated with the desired purified alcohol (isopropyl or secondary butyl) at 60" to 80°C to form the dithiophosphoric acid derivative. Purification was effected in the same manner as for the nonradioactive material. The effectiveness of the purification method was established by the isotopic dilution method." The same technique was used to show that only negligible decomposition of neutral or of carbonate solutions of the dithiophosphate took place over a period of a day; this was true whether or not galena was suspended in the solutions. Procedures: Radioactivity: The activity of a solution was obtained by counting with a small glass-jacketed, silvered Geiger counter, using a conventional scaling circuit.&apos; Crystals and other solids were counted under a bell-shaped, mica-window counter tube. Sufficient counts were made so that the probable error of counting was of the order of +I-3 pct. The specific activity in terms of counts per minute

Jan 1, 1951

By F. Weinberg, E. H. McLaren

Dilute binary alloys have been solidified under controlled thermal conditions, and solute distributions, temperatures during freezing and melting, and the position and morphology of the solid-liquid interfaces have been examined. Binary Zn, Sn, and Pb based alloys were investigated with solute additions of AgU°, AU°, Sb&apos;24, TlZo4, snH3, and Zns5. Approximately 150 cc of metal was used in each ingot. Appreciable solute segregation occurred in most of the ingots, depending on the solute concentrations, distribution coefficients. and freezing conditions of the alloys. Measurements were made of the general solute distribution by sectioning layers of material from the ingot circumference and counting the activity of each layer. Local segregation was detected by autoradiographic techniques. Freezing and melting curves for the same alloys were obtained by precision resistance thermometry. These curves are related to the corresponding measured solute distribution curves and those calculated from theoretical expressions. The position and morphology of solid-liquid interfaces during freezing were determined by both decanting and quenching techniques. The results indicate the manner in which freezing and remelting occurred under the conditions investigated. If an isomorphous dilute binary alloy is solidified or melted under equilibrium conditions, the concentration of solute in both solid and liquid would be uniform at all stages in the freezing process. The solute concentrations during freezing could be determined from the phase diagram of the alloy, and the freezing curves would have an alloy slope defined by the solidus and liquidus points. In practice, however, solidification under equilibrium conditions rarely, if ever, occurs, primarily because of low rates of diffusion in the solid and incomplete mixing in the liquid. As a result, the solute distribution and the shape of the freezing and melting curves obtained in a given system depends on a number of factors. These include, in addition to the extent of diffusion and mixing, the initial solute distribution and concentration, the distribution coefficient, the rate of freezing or melting, supercooling of the liquid, the morphology of the solid-liquid interface, and the effect of annealing. The purpose of the present investigation was to determine, experimentally, actual solute distributions and thermal curves for a series of dilute binary alloys, using tracer techniques to determine the solute distributions and precision resistance thermometry for the temperature measurements. In addition, observations were made of the positions and morphologies of the solid-liquid interfaces during freezing and remelting. The distribution of solute in a dilute alloy, assuming complete mixing in the liquid, and negligible diffusion in the solid, is given by1 where Co is the mean solute concentration of the alloy, C, is the solute concentration in the solidg is the fraction of the ingot solidified, and ko is the equilibrium solute distribution coefficient (assumed to be independent of concentration for dilute alloys). For the case of partial mixing, ko can be replaced by an effective distribution coefficient k. If solute transport in the liquid is assumed to be entirely due to diffusion, with negligible diffusion in the solid, then the solute distribution in the solid is given by Tiller et a1.&apos; as where R is the rate of advance of the solid-liquid interface, D is the solute diffusion coefficient in the liquid, and x is the position of the interface from the start of freezing. The temperature at the interface TI is related to the solute concentration of the solid at the interface C, by the expression where Tm is the freezing temperature of the pure solvent and ml is the slope of the liquidus for the binary alloy being considered. Accordingly, by relating TL to the actual temperatures measured in the present experiments, solute distributions can be deduced from the temperature measurements. Using the above expressions, the measured solute distribution of a given alloy may be compared to the theoretical distributions, as well as to the distribution calculated from the observed freezing curve. EXPERIMENTAL The solvent materials used in making up the binary alloys were C. P. New Jersey zinc (99.999 pct), Vulcan Extra Pure Tin (99.999 pet), and Tadanac lead (99.998 pct). The solute elements were of 99.99 pct purity or better, and were activated in the

Jan 1, 1963

By Charles C. Peterson, Ronald E. Enstrom

Significant improvements have been made in the ursine systern for epitaxial vapor gvowtlz of Gds. The electron concentration has been reduced to below 1015 cm-3 with electron-mobility values as high as 7200 sq on per v-sec at 30PK and 60,000 sq cw7 per v-sec at 77°K. These high-purity layers have been successfully gvown on n&apos; layers with thicknesses as snzall as 1 p and as large as 125 p. Furthermore, high-purity layers have been incorporated into an structure without degradation of electrical properties. Different impurity gradients at the n&apos; to n interface were achieved. Using these structures, Gunn oscillators have been operated at frequencies as low as 0.4 GHz and as high as 40 GHz, which equals the highest valzle yet reported. The frequencies nleasured are in reasonable apeenlent with those calculated fron7 the n-layer widths and the transit times. CW operatiolz has been achieved from 3 to 20 GHz. In Gunn devices for >40 GHz operation (i.e., witk an n-layer tkickness below 2.5p), the n layer has been generally too pure to permit oscillations to occuv. ThE mechanism of the Gunn oscillator&apos; is now generally understood.2-5 Electron transfer from the (000) minimum to the ( 100) minima leads to a voltage-controlled negative resistance, which results in the formation of domains of high and low electric fields which travel through the crystal with a characteristic velocity of about 107 cm per sec. When a constant voltage is applied to the sample, this domain motion generaliy leads to oscillations at the transit time frequency, fT = 107/L, where L is the thickness of the active n layer.&apos; In this case, a single high-field domain is nucleated at the cathode and grows until it reaches the anode, where it is absorbed; a new domain then forms at the cathode.5 There are related modes of oscillation of the same device in which a tuned circuit causes a periodic variation of the voltage across the device. The frequency then may be either above or below fT. It has been shown that oscillations will occur when n • L 2 1012 cm"2, where n is the carrier concentration in the active region. At lower doping densities, 10" 5 n . L 5 lo&apos;&apos;, the doping is said to be subcritical for oscillation and the description of the device operation changes since the domains become comparable in thickness to the device thickness, L. Nevertheless, such devices have been found to be "active" and, in fact, are useful as amplifiers.778 For CW operation, one wishes to minimize the power dissipation, which is related to the electric field, E, and the low-field resistivity, p. Below threshold, the power dissipation density is approximately Typically, at threshold where E -- 3 x 103 v per cm, the dissipation density is 107/p watts per cu cm, p in ohm-cm. Furthermore, it is desirable to provide good conduction of the heat generated in the active region. The temperature rise, AT, across the active layer of a sandwich structure, with a heat sink on one side, is given by AT = (E&apos;/P)(L~/~~) where it is assumed that the thermal conductivity, u, and resistivity, p, of GaAs do not vary appreciably over the range of temperature and electric fields existing in the layer. This expression shows that, subject to maintaining an appropriate (n . L) product, CW operation is favored by a higher resistivity, a lower threshold field, and a thinner device. The above device requirements place some severe restrictions on the materials to be used, especially for high-frequency (>10 GHz) operation. Here, it is necessary to prepare very thin (<10 µ) layers of GaAs, in a state of high purity (-1 ohm-cm), complete with very low-resistance ohmic contacts. Problems of handling such thin and fragile samples become inereasingly severe as the frequency is raised. Furthermore, while good low-resistance ohmic contacts are easily applied to low-resistivity GaAs, contacting difficulties are encountered for these higher-resistivity materials.

Jan 1, 1968

By M. V. Lowry

This paper outlines the economic considerations that led to the recent installation of a thickener at St. Joseph LeadCo.&apos;s Balmat, N.Y. mill. To incorporate storage of concentrates, they decided upon a single, unusually deep tank to be used for both storage and thickening. The abnormally high rake lift required was achieved by the novel feature of having the entire drive platform raise and lower. Initial operating problems are described along with the satisfactory solutions. Total costs for concentrate handling before and after installation of the storage thickeners are compared. In Northern New York State, the St. Joseph Lead Co. conducts a 2100 tpd zinc mining and milling operation at its Balmat Plant. Conventional crushing, grinding, and flotation is used to produce a zinc concentrate. For 32 years, from 1930 through 1962, the plant operated with the concentrates going directly from flotation to the filtering, drying, and loading facilities. All of the concentrate handling equipment, therefore, had to operate 24 hrs per day. It also had to accommodate wide fluctuations in tonnage, due to variations in the grade of ore. Many problems were thereby encountered in filtering and drying, but it was never concluded that the cost could be justified for installing a conventional, large diameter thickener. Finally, in 1963, they installed an unconventional thickener that is unusual in several respects. The tank depth is 16 ft instead of the normal 8 ft. The rakes are raised and lowered by having the main drive gear located on a movable platform. Automatic controls are programmed to raise the drive platform and rakes during one shift of storage. During the next two shifts of withdrawal, the controls are programmed to lower the drive platform and rakes. After three years of operating this unique storage-thickener, the conclusion is that a correct choice was made from several alternative schemes for storing concentrates. The various considerations involved in the final selection will here be analyzed, and the design features and installation details will be de- scribed. Operating problems and modifications made to the equipment will next be considered, and finally the effects upon filtering, drying, and loading will be reported. SELECTION CONSIDERATIONS Even without a thickener, concentrate handling had become reasonably efficient at the Balmat Mill by the year 1962. A tonnage that varied between 6 and 25 tph was being handled fairly satisfactorily. The filter cake moisture of 8%% seemed quite respectable, except that two large filters were usually in service. Three filters often had to be operated in order to prevent excessive filter overflow with consequent loss of values. This situation occurred when zinc concentrates from flotation exceeded 20 tph and when filter cloth blinding developed. Over the years, numerous small improvements had been effected, such as: selection of a filter cloth with better flow characteristics; addition of flocculant to the filter feed; and, in 1953, the smallest of three drum filters was replaced by a five disc, 8&apos; 10" diam filter with a bottom agitator and snap blow. As another example of fairly efficient operation, the oil-fired dryer was doing a seemingly satisfactory job of further reducing the moisture content to an average of 3%. However, there were often considerable dust losses when the tonnage suddenly dropped off to 6 tph. At high tonnage rates, the resulting high moistures of 4% and 5% caused difficulties due to concentrates sticking and freezing in chutes leading to the box car loaders. (Drying of concentrates is justified because of the high freight rate for the long shipping distance to the company owned smelter at Josephtown, Pa. near Pittsburgh.) It was long realized that installation of a thickener would permit the subsequent steps of concentrate handling to proceed more smoothly and would alleviate the problems existing in the filtering, drying, and loading operations. It was also realized that even a slightly lower filter cake moisture could mean considerable savings in fuel oil. However, all these factors were never quite sufficient to warrant the expense of a thickener until three more considerations were added to the picture. First, the advent of flocculants meant that a thickener of relatively small area could now be considered. Secondly, a complete renovation of the zinc flotation circuit was made between 1957 and 1961. The

Jan 1, 1967

By Ronald K. Linde

Metastable Ag-Cu solid solutions at two compositions beyond the maximum limits of solubility obtainable by quenching from the solid state were obtained by rapid quenching from the liquid state. The process of transformation of the metastable solid solutions into the equilibrium phases was studied by means of X-ray diffraction intensity measurements and electrical-resistance measurements; metallographic examination was also performed. Results are compared with those for an alloy quenched from the solid state at a composition which exists as a solid solution at elevated temperatures. Activation energies for the trans-formation were computed to be 33,600 * 2500 and 33,100 * 2000 cal per g mole for the 60.1 and 75.0 at. pct Ag;Cu compositions, respectively THE kinetics of precipitation in supersaturated solid solutions obtained by quenching from high temperature has been thoroughly studied in many alloys, and detailed investigations of the early stages of nucleation and subsequent growth of the equilibrium phases have been carried out. All of these studies have been made on alloys quenched from the solid state, however, and it is the purpose of this paper to present some of the results on the kinetics of transformation of metastable solid solutions in Ag-Cu binary alloys whose compositions lie outside of the solid-state limits of solubility. Limits of solubility in the solid state can be increased by rapid quenching from the liquid state. This was first demonstrated by Falkenhagen and Hofmann, who obtained appreciable increases in solubility limits of the elements of the first transition series in aluminum by rapid cooling of liquid alloys through contact with a copper mold maintained at liquid nitrogen temperature.&apos; The same basic idea of cooling a liquid alloy by conduction onto a substrate rather than by convection of a gas or liquid (conventional quenching) has been recently revived and extended to very high rates of cooling by Duwez and his collaborators. The new technique was first applied to Ag-Cu alloys because the existence of a eutectic in this system is not predicted by the Hume-Rothery rules of alloy phase formation. Rapid quenching from the liquid state resulted in metastable solid solutions at all concentrations. The kinetics of decomposition of these metastable solid solutions containing 60.1 at. pct Ag;Cu (eutectic composition) and 75.0 at. pct Ag;Cu is the subject of the present paper. EXPERIMENTAL TECHNIQUES Alloys were made in 10- to 20-g heats using silver of purity 99.99 pet and copper of purity >99.999 pct. Components were weighed accurately to 0.1 mg and melted by induction heating in an alumina crucible under an atmosphere of hydrogen. Weight losses amounted to less than 0.02 ~ct in all cases. Rapid cooling from the liquid state was achieved in an apparatus of the type described by Pietrokowsky. 3 In this technique a freely falling drop of molten alloy is caught between a stationary anvil and a fast moving piston, both anvil and piston being lined with copper. The resulting foils were approximately circular (about 2 cm in diam) and from about 50 to 150 u in thickness, depending on the pressure driving the piston. Metallographic samples were prepared for polishing by edge-mounting foils in Quickmount. The mostef-fective etching procedure was found to be: 10 parts NH4OH (concentrated), 1 part H2O2 (30 pct), 5 parts H2O, swab for 5 to 10 sec. The transformation of the metastable alloys to equilibrium was followed by X-ray diffraction and electrical-resistivity measurements. The peak intensity and the integrated intensity of suitable Bragg reflections of the metastable and stable solid solutions were studied with a Norelco dif-fractometer, using nickel-filtered copper KO radiation. The peaks were step-scanned at intervals of 0.05 deg. Using either a Kelvin bridge or a null circuit with a Leeds and Northrup K-2 potentiometer, the changes in electrical resistance due to phase transformation were measured for specimens cut into strips about 20 mm long and 4 mm wide. All resistivity measurements were carried out at liquid-nitrogen temperature. The isothermal heat treatment was accomplished by sealing the samples in argon-filled aluminum containers with stainless-steel O-rings. Containers were then suspended in wax baths with temperatures in the

Jan 1, 1967

By R. I. Jaffee, F. C. Holden, H. R. Ogden

ECENT papers dealing with the properties of unalloyed iodide titanium have been directed primarily at the determination of base-line properties for alloy investigations. Early work was limited to a few tests because of the limited availability of iodide titanium at the time. In the results of papers by Campbell et al.,1 Gonser and Litton,2 Jaffee and Campbell,3 inlay and Snyder,4 and Jaffee, Ogden, and Maykuth, data on mechanical properties are presented for unalloyed iodide titanium in the annealed and cold-worked conditions. Data are presented in this paper which show the effects of heat treatment on the structure and mechanical properties of commercially produced iodide titanium. Correlation is made between microstruc-tural variables and the mechanical properties. Experimental Procedures Melting Stock: The melting stock used was as-deposited iodide titanium, produced by New Jersey Zinc Co. The furnished analysis showed the following range of impurities: N, 0.004 to 0.008 pct; Mn, 0.005 to 0.013; Fe, 0.0035 to 0.025; Al, 0.013 to 0.015; Mo, 0.0015; Pb, 0.0045 to 0.0065; Cu, 0.0015 to 0.002; Sn, 0.001 to 0.01; Mg, 0.0015 to 0.002; and Ni, 0.003. Hydrogen content as determined by vacuum-fusion analysis was 0.0091 wt pct (0.44 atomic pct) after arc melting and fabrication. Nitrogen analysis on the arc-melted and fabricated titanium showed a content of less than 0.002 pct N. The average hardness of the furnished stock was Rf 70, or approximately 85 VHN. Melting Procedure: The as-deposited rods were rolled, sheared, and degreased in preparation for arc melting. The charge was arc-melted with a tungsten electrode in a water-cooled copper crucible under a positive pressure of high purity (99.96 pct) argon. The final ingot was approximately 2 in. in diameter and showed no increase in hardness over that of the initial stock. Fabrication: Heating for fabrication was done in air. It was begun by forging the ingot into a 3/4 in. diam rod, at an initial temperature of 1600°F. Scale was removed by sandblasting. The rod was then swaged to 1/4 in. diam at room temperature through a series of 20 dies, with approximately 10 pct reduction in area between each die. An anneal of 1 hr at 850 °C in air was given after the 1/2 in. die, such that the final cold reduction was 75 pct. Sections cut from this rod were used for test and microstructure specimens. Heat Treatment: Heat treatments were carried out in resistance tube furnaces with stainless-steel linings, under an atmosphere of gettered argon. As further protection against contamination, the specimens were packed in titanium turnings in a titanium sleeve. Control experiments have shown negligible hardness increases with this method, indicating that contamination from oxygen and nitrogen is slight. Three cooling rates were employed in this work; these have been designated as water quenching, argon cooling (to simulate air cooling under a controlled atmosphere), and furnace cooling. The cooling rate for an argon cool is 100°C per min for the first minute, with an average cooling rate of 35°C per min over a 15-min period. A furnace cool requires about 10 hr, with an average cooling rate of 3.6oC per min during the first hour, and an average cooling rate of 1.2°C per min over the 10-hr period. Microimpact Test: The specimen adopted was based on the cylindrical Izod Type Y specimen (ASTM, E23-41T). All dimensions were reduced to half scale, including the notch radius. Specifications are shown in Fig. 1. The specimen is held vertically in an adapter and broken as a cantilever beam. Impact tests were run on a constant-velocity (11.34 ft per sec) Tinius Olsen impact testing machine with a total available energy of 100 in.-lb. Tests were made to determine the correlation between this microimpact and the standard V-notch Charpy impact test. Curves showing impact energy as a function of temperature for both impact tests are plotted in Fig. 2. Transition temperatures, when they occur, are about the same for both impact tests. All three titanium-rich materials have the same conversion factor, 10. Tensile Testing: Tensile tests were conducted on Baldwin-Southwark testing machines using the 600, 2400, or 3000 1b range. Specifications for the test specimen were taken from the 1948 edition of the ASM Metals Handbook, and are shown in Fig. 1. Strain measurements were made using an SR-4 resistance gage (Type A-7) cemented to the reduced section in conjunction with a lever-type extenso-meter. Readings on the SR-4 strain indicator were

Jan 1, 1954

By D. C. Hilty

It has long been known that in melting high-chromium steels, some of the carbon might be oxidized out of the melt without excessive simultaneous oxidation of chromium, and that higher temperatures favor retention of chromium. The advent of oxygen injection as a tool for rapid decarburization of a steel bath permits significantly higher bath temperatures, and it was quickly recognized that the use of oxygen injection facilitated the oxidation of carbon to low levels in the presence of relatively high residual chromium contents. Up to the present time, however, specific data pertaining to the chro-mium-carbon-temperature relations in chromium steel refining have not been available. Individual steelmakers have evolved practices more or less empirically, but there has been very little real basis for predicting how effective any given practice can be in permitting maximum oxidation of carbon with minimum loss of chromium. The current investigation, therefore, was undertaken in an effort to establish the fundamental carbon-chromium relationship in molten iron under oxidizing conditions. As reported below, the equilibrium constant and the influence of temperature on that constant have been derived for the iron-chromium-carbon-oxygen reaction in the range of chromium steel compositions with what appears to be a fair degree of precision. The practical application of the result will be obvious. Experimental Procedure The laboratory investigation was carried out on chromium steel heats melted in a magnesia crucible in a 100-lb capacity induction furnace at the Union Carbide and Carbon Re- search Laboratories. The charges for the heats consisted of Armco iron, low-carbon chromium metal, and high-carbon chromium metal, the relative proportions of which were calculated so that the various heats would contain from approximately 0.06 pct carbon and 8 pct chromium to 0.40 pct carbon and 30 pct chromium at melt-down. When the charges were melted, the bath temperatures were raised to the desired level, and the heats were then decarburized by successive injections of oxygen at the slag-metal interface through a ½-in. diam silica tube at a pressure of 30 psi. The duration of the oxygen injections was from 30 sec to 2 min. at intervals of approximately 5 to 30 min. It did not appear that length or frequency of the injection periods had any significant effect on the results; cansequently, no effort was made to hold them constant and they were controlled only as was expedient to the general working of the heats. Between successive injections, the heats were sampled by means of a copper suction-tube sampler that yields a sound, rapidly-solidified sample representative of the composition of the molten metal at the temperature of sampling. This sampling device is a modification of the one described by Taylor and Chipman.1 An attempt was made to vary bath temperatures between samples, but it quickly became evident that, unless the variations were small or unless the new temperature was maintained for a minimum of 15 min. during which an injection of oxygen was made in order to accelerate the reactions, a very wide departure from equilibrium resulted. For most of the runs, therefore, temperature was maintained relatively constant at approximately 1750 or 1820°C. A few reliable observations at other temperatures, however, were obtained. Temperature Measurement The high temperatures involved in this investigation were measured by the radiation method, utilizing a Ray-O-Tube focused on the closed end of a refractory tube immersed in the metal bath. The immersion tubes employed were high-purity alumina tubes specially prepared by the Tona-wanda Laboratory of The Linde Air Products Co. These tubes were quite sturdy under reasonable mechanical stress at high temperature. They were unusually resistant to thermal shock, and chemical attack on them by the melts was slow. With care, it was found possible to keep these tubes continuously immersed in a heat for as long as 5 hr at temperatures up to 1850°C, before failure by fluxing occurred. The Ray-O-Tube—alumina tube assemblage was similar to those supplied commercially for lower temperature applications. In operation, the alumina tube was slowly immersed in the molten metal to a depth of approximately 5 in., and the device was then clamped solidly to a supporting jig where it remained for the duration of the run. A photograph of the equipment, in operation with Ray-O-Tube in place and oxygen injection in progress, is shown in Fig 1. When in position in a heat, the instrument was calibrated by means of an immersion thermocouple and an optical pyrometer. For calibration through the range of temperatures from 1500 to 1650°C, a platinum -platinum + 10 pct rhodium thermocouple in a silica tube was immersed alongside the alumina tube. Output of the Ray-O-Tube in millivolts and the

Jan 1, 1950

By W. C. Clarke

An empirical study of the nature of the embrittle-ment which occurs in martensitic and semiaustenitic precipitation hardening stainless steels upon exposure at temperatures of from about 550" to 875°F has been undertaken. This work was aimed at determining cazusation and means of controlling this phenomenon. A commercial copper bearing precipitation hardening alloy was used as a basic material for study. The effect of heat treatment variables was studied as was the effect of variations in analysis. It is concluded from the evidence that martensitic and semiaustenitic precipitation hardening stainless steels are subject to 885°F ernbrittlement similar to that observed in the straight chromium stainless steels. A characteristic of precipitation hardening stainless steels which has limited their use in certain applications is that they embrittle when held at temperatures in the range of from about 550" to 900°F. This is true to a more or less degree in all currently available alloys, either the basically martensitic type or the semiaustenitic type. The rate of embrittlement varies markedly with exposure temperature, being low at 550" to 600°F and in-creasing as the temperature increases. In spite of this embrittlement, these alloys with their unique combination of high mechanical strength, reasonable toughness, and good corrosion resistance are used in many hundreds of applications. Nevertheless, there are potential applications where the embrittle-ment described is a limiting factor. The purpose of this investigation was to study the embrittlement of these alloys and to find a way to control or eliminate it. GUIDELINES USED IN PRESENT WORK The work reported in this paper is based on a study of 17-4 PH*, a very widely used alloy. It has been used *Trademark of Armco Steel Corp., licensor. in pressurized water and boiling water atomic reactors operating at about 550°F for a number of years. As the life of such equipment is extended or operating temperatures are raised, the possibility of embrit-tlement becomes of increasing concern to materials engineers. Much investigation work was done with respect to the use of 17-4 PH at 550°F. K. C. Antony&apos; states "Such estimation" (of an activation energy for the diffusion of chromium in iron) "would indicate W. C. CLARKE, Jr. is Senior Research Engineer, Advanced Materials Division, Armco Steel Corp., Baltimore. Md. This manuscript is based on a talk presented at the symposium on New Developments in Stainless Steel, sponsored by the IMD Corrosion Resistant Metals Committee, Detroit, Mich., October 14-15, 1968. significant secondary aging is improbable at temperatures less than 700°F within normal component service life". This statement is modified however by the recognition of the accelerating tendencies of applied stress and internal stress as well as the possible effect of irradiation. In this investigation of 17-4 PH, the H 1100 (1900°F-1 hr oil quench or air cool + 1100°F-4 hr-air cool) condition was used, partly because this condition is normally used in atomic reactors. As shown later, the precipitation hardening temperature has no real bearing on the rate of embrittlement. An exposure temperature of 800°F was selected since embrittlement at 800°F is rapid, permitting development of relative data in time periods of 400 to 500 hr. For those not familiar, a nominal present day analysis of 17-4 PH is: C Mn P S SiCrNiCuCbN 0.04 0.30 0.015 0.015 0.60 16.00 4.30 3.25 0.23 0.030 TYPICAL BEHAVIOR OF 17-4 PH Figs. 1 and 2 show the behavior of a commercial heat of 17-4 PH under the conditions defined. Characteristically, 800°F exposure causes a rapid drop in Charpy V-notch impact strength. Tensile and yield strengths gradually increase and a gradual loss of elongation and reduction of area occur, accompanied by an increase in hardness. Notched tensile strength increases to 125 hr exposure and then sharply decreases after exposure for 500 hr. The notched vs un-notched tensile ratio remains virtually constant to exposure for 125 hr (1.67 to 1.56) but drops to 1.15 after 500 hr. Tensile ductility is not alarmingly affected even after much longer exposure times than these. For example, samples aged at 1100°F for 1 hr exposed at 800°F for 4000 hr showed a drop in elonga-tion of from 13 to 11 pct and in reduction of area of from 58 to 37 pct. Notched impact is the property SmIgth KSI 3JJ[/__________^UTS-Nrteh 260 ;/- 240 —^ 200 UTS-Smooth________._, o ioo mo Sob" m Too Hours Exposure Fig. 1—Effect of exposure at 800°F for various times on notched tensile strength and smooth tensile and 0.2 pct yield strength of 17-4 PH in the H 1100 condition.

Jan 1, 1970