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Prof. Henry M. Howe, New York, N. Y. (communication to the Secretary*):—The results presented in Mr. Mathewson's paper are of very great interest. Perhaps the most interesting is that shown in Table I. and Fig. 1, in which the arsenic decreases in 10 min. from 0.22 to 0.07 per cent. This is such a large decrease that it is to be hoped that Mr. Mathewson will verify it by repeating his tests with a matte of similar composition. This is particularly to be desired, because the simultaneous removal of sulphur appears to be very small. All three blows further agree in showing that the removal of arsenic is in the very early part of the blow. The fact that the removal of arsenic then ceases abruptly cannot be explained completely by the fact that the arsenic has fallen so low as by its lowness to resist further removal, because in the first blow the arsenic becomes stationary at 0.06 per cent., whereas in the third blow it does not become stationary until it has fallen to 0.034 per cent. The abrupt ending of the removal of arsenic is more likely to find its explanation in the composition of the slag, and it is to be hoped that Mr. Mathewson will supplement these extremely valuable data with others in which the composition of the slag at the same time shall be given. The sudden rise of arsenic toward the end of the third blow I suppose must be referred to an error in sampling or analysis.

Jan 1, 1908

By William M. Agar

Now that both the national and state legislatures are seeking ways of reducing expenses, the appropriations for geological investigation and for the study of mineral resources have been greatly reduced and in some cases entirely stopped. Is this policy justified? The sums saved by these reductions are always small. If the work accomplished by these surveys has no intrinsic value to the community, but is only an interesting way of spending some of the people's money when that money is plentiful, there should be no surveys at any time. If, on the other hand, they perform a necessary function, it is indeed foolish and short-sighted to save these paltry sums to the detriment of the community. It is not my purpose to undertake a general justification of geological surveys. Rather, 1 wish to call attention to the work of one state survey that has been discontinued, and to indicate ways in which the work was valuable to the people, even though that state's mineral resources are not very important, and why it should be continued. Though at one time in Connecticut's history iron mining and smelting were important, and copper was mined in appreciable quantities in two localities, it is many years since it could be regarded as a mining state. At present Connecticut stands far down among the states in the value of its mineral products. Granite for architectural and structural purposes, trap rock, clay, agricultural lime, feldspar and quartz, together with a few other nonmetallic minerals, are produced in appreciable quantities, but the natural resources of the state consist at present much more largely in its attractive scenery, woods, lakes, shore line, and in the opportunities that these offer for rest and recreation. Is the prosecution of geological work in such a state a waste of the taxpayer's money? Or can it be that though it was formerly valuable, the work is now completed? The only way to answer these questions is to review the geological work done in the past and to consider the type of information the people want, as reflected in the requests that have been received by the director of the survey in the last few years. There have been two periods during which Connecticut has supported geological work within its borders. The first began in 1835 when Percival

Jan 1, 1935

By William M. Agar

Now that both the national and state legislatures are seeking ways of reducing expenses, the appropriations for geological investigation and for the study of mineral resources have been greatly reduced and in some cases entirely stopped. Is this policy justified? The sums saved by these reductions are always small. If the work accomplished by these surveys has no intrinsic value to the community, but is only an interesting way of spending some of the people's money when that money is plentiful, there should be no surveys at any time. If, on the other hand, they perform a necessary function, it is indeed foolish and short-sighted to save these paltry sums to the detriment of the community. It is not my purpose to undertake a general justification of geological surveys. Rather, 1 wish to call attention to the work of one state survey that has been discontinued, and to indicate ways in which the work was valuable to the people, even though that state's mineral resources are not very important, and why it should be continued. Though at one time in Connecticut's history iron mining and smelting were important, and copper was mined in appreciable quantities in two localities, it is many years since it could be regarded as a mining state. At present Connecticut stands far down among the states in the value of its mineral products. Granite for architectural and structural purposes, trap rock, clay, agricultural lime, feldspar and quartz, together with a few other nonmetallic minerals, are produced in appreciable quantities, but the natural resources of the state consist at present much more largely in its attractive scenery, woods, lakes, shore line, and in the opportunities that these offer for rest and recreation. Is the prosecution of geological work in such a state a waste of the taxpayer's money? Or can it be that though it was formerly valuable, the work is now completed? The only way to answer these questions is to review the geological work done in the past and to consider the type of information the people want, as reflected in the requests that have been received by the director of the survey in the last few years. There have been two periods during which Connecticut has supported geological work within its borders. The first began in 1835 when Percival

Jan 1, 1935

By William Kent

About a year ago, the writer had occasion to assist Mr. John L. Gill, Jr., of the Pittsburgh Car-wheel Works, in malting a trial of his new testing machine. A piece of Bessemer steel, of about .34 carbon, was used as a test-piece, turned to a sectional area of one square inch for a length of five inches between shoulders, and provided with large round heads, which were fitted to what is called the " Wade grip." Measurements of elongation, accurate to 1/10000 of an inch, were made by a special apparatus containing two micrometer screws, with electric contact attachment. The machine was operated by hand-power. Readings of elongation were taken at each increment of 2000 pounds stress per square inch, up to 44,000 pounds, the elongation corresponding to the latter figure being .0085 inch. The elongations up to this point remained almost exactly proportional to the stresses, the strain diagram, as shown on the Plate, from the origin to the point A, being a straight line (the deviation of no observation from the straight line being greater than .0002 inch). There being no indication of an approach to the elastic limit, the attendant at the hand-wheel and scale-beam wits told to proceed as usual and move the poise forward 2000 pounds, or to 46,000, and turn the wheel until the beam was balanced. Noticing that the beam did not rise in the usual time, and that the elongation was increasing rapidly, the writer called to the attendant to stop turning the hantl-wheel and to push the poise backwards until the scale-beam balanced. Instead of balancing at between 44,000 and 46,000 pounds, as was expected, it balanced at 40,000 pounds, and the elongation had increased to .0343 inch (or more than four times that recorded at 34,000 pounds), as plotted at the point B, the dotted line between A and B showing the supposed course of' the strain diagram between these two points. Immediately upon recording the elongation the writer said to Mr. Gill, who was standing alongside: "This phenomena is without a parallel in my experience, and I know of but one way to account for it. The piece. must actually be broken in its interior, through a flaw or weak portion of the metal. The test-piece is now equivalent to a tube, with

Jan 1, 1880

By E. V. Foran

AMONG the many newer disclosures that have accompanied the petroleum industry's progressively deeper exploratory drilling is the increased frequency with which the operators are encountering reservoirs of which the products at the time of discovery exist wholly or predomi-nantly in the gas phase. The popular term "distillate pool" is most often used within the industry when referring to this type of reservoir, and in this paper the term is used in that sense. These pools have been observed in many different surroundings. In some places the entire reservoir is filled with the wet gas containing its water-white condensible products in the gas phase but apparently with no dark oil or other hydrocarbon liquids in the reservoir. The 8200-ft. horizon in the Big Lake field in West Texas is a good example of this type. At other places the wet, or distillate-bearing, gas may form a large gas cap directly overlying and in contact with an extensive reserve of dark oil. Oklahoma City, Kettleman Hills, the 8500-ft. horizon in the Big Lake field, and many others are examples of this type. There are also many places in which the wet distillate-bearing gas fills nearly the entire reservoir space, while the dark oil occupies only a narrow, bandlike zone, which marks the boundary of the reservoir along its lowest productive contour. The heavier components of these deep pool gases are generally water white in color, a large portion of which are condensible in a conventional oil and gas separator, when operated under pressures of the order of 400 to 800 lb. per sq. in. Although the water-white products recovered in the separator sometimes have end points of 650° F. and higher on standard Engler distillation, they occur only in the gas phase in the reservoir. It is entirely possible that even some small amounts of dark colored reservoir components may occur in the gas phase when the temperature is 300° F. and pressures exceed 6000 lb. per sq. in. Close approaches to these conditions have been encountered in the deepest producing wells. The condensible content of the deep reservoir gases of the many pools that have been tested or produced to date shows a wide variation in both condensible content, per 1000 cu. ft. of gas, and physical character.

Jan 1, 1939

By William N. Poundstone

To forecast the future of coal intelligently, one must first consider the market outlook. Coal can be used for many purposes, but currently about 25% of it goes into metallurgical coke, while 75% serves as fuel. Coal for metallurgical coke should show a modest growth for at least the next decade, primarily because of increased exports, but the important factor in the coal outlook is energy. Today, this country faces a serious shortage of energy, and its requirements are growing at an almost unbelievable rate. Last year the United States consumed 68 quadrillion Btu, or double the energy it consumed in 1950, and, if projections are correct, this figure will grow to 150 quadrillion Btu by 1990. Thus, in one working lifetime, the nation's energy consumption is expected to increase by roughly five times. The future of coal hinges on its degree of participation in meeting this energy need. The Federal Power Commission predicts that coal consumed by electric utilities will increase about 4-576 per yr, reaching 472 million tons by 1980 and 615 million tons by 1990. If this projection proves correct, 1980 consumption will be double that of last year and, even more startling, the utility industry alone will use more coal in 1990 than the entire coal industry produced in 1970. This is a tremendous increase, but in view of the projected energy growth, it is difficult to see how this country's needs can be satisfied with- out a coal contribution of this magnitude. Unless some new source such as solar energy is discovered, it must be assumed that the projected energy will come principally from the present-day sources-oil, gas, nuclear materials and coal.

Jan 1, 1971

By C. S. Smith

The partial persistence of grain size and grain shape on melting and resolidifying crystalline substances, as well as the general effects of pre-solidifi-cation and of superheating on nuclea-tion rate, have been attributed to the presence in the liquid of crystalline impurities bearing some structural relation to the principal solid and hence capable of serving as nuclei for its crystallization. That such lattice matching does sometimes occur is well established, but it is surely improbable that a compound possessing both high-temperature stability and the requisite structural similarity should be naturally available in almost every system. It is now suggested that a mechanism for producing appropriate nuclei exists in the crystallization of all but absolutely pure substances. Impurities can he of many kinds. Some will give rise to compounds that are less soluble in the solid crystal than in the liquid. A . crystal forming from a melt or solution containing these will be supersaturated in some degree, and the foreign atoms will commence to segregate during or after solidification. They do not necessarily form pure, unstrained, crystals of the precipitating compound with its normal composition and structure. Generally—at least in the metallurgical examples that have been most studied —the first recognizable nuclei have lattice coherency with the surrounding matrix and have practically the same crystal structure and lattice spacing as the material in which they form, differ- ing only in composition. An extremely thin oxide layer formed under appropriate conditions on or in a solid metal will tend to be coherent with the metal lattice, and the first few atom layers of the oxide in contact with the metal will differ in both composition and structure from massive crystalliue oxide. It will resemble the metal more closely than the oxide. If it did not subsequently change on melting, such an oxide film would constitute an almost perfect "template" or two-dimensional matrix to hold a few atoms of the primary metal in the exact array needed to form a nucleus for solidification. Any precipitating material will adopt the correct spacing over a few atom diameters, and with appropriate substitution of other atoms of the correct sizes a precipitate can match the parent lattice over large areas of cohererlce without excessive strain. In metals, the componentsof a host of oxides, silicides, borides, nitrides, carbides, sulphides, and other compounds are always present in minute amounts. Some of these will be eliminated because of extreme insolubility in the liquid, while others will remain soluble in the solid; occasionally, however, particles will be formed in the solid by coherent precipitation—not of pure substances, but of whatever assortment of atoms best satisfies the joint requirements of availability, affinity, and appropriate average atom size. Of these, many will disappear on melting, but a few will persist unmelted as little rafts of complex composition, maintaining in the liquid a surface of the exact geometry needed to nucleate the solid on subsequent cooling. Their number and size distribution will determine the resulting grain size; the dimensions of the largest will determine the degree of undercooling that can occur before solidification commences. A similar process may operate in the case of phase transformations in solids. The coherent precipitation of supersaturated minor constituents in a low-temperature phase should facilitate the nucleation of this phase on subsequent cooling after transforming into another phase at higher temperatures. At first sight this hypothesis may seem little different from the older ones. It should therefore be emphasized that it does not depend on chance to provide the right compound to form a nucleus, but postulates that every crystal, containing minute amourlts of any of a wide range of impurities, automatically engenders particles having the correct surface structure to serve as nuclei for subsequent solidification. It could operate in any system, metallic, organic, or inorganic. The suggested mechanism is closely analogous to the production of antibodies from antigens in living organisms—indeed, it was a description of the matrix theory of this process that suggested to the writer its crystalline analogue.

Jan 1, 1950

By T. A. Wright

THE-Institute of Metals Division opened on Tuesday afternoon with Wheeler P. Davey as chairman and G. E. Edmunds as vice-chairman. Four papers were on the program, two being of a fundamental character. A paper on "Internal Surface in Minerals and Metals," by R. S. Dean and J. Roster, was presented by the latter; C. H. Gottschalk discussed coercive forces in powders. In answer to a question by the chairman. The author stated that he had not investigated [he temperature coefficient of resistance. A paper on "Recrystallization and Grain Growth in Cold Worked Polycrystalline Metals," by L. W. Eastwood, Arthur E. Bousu, 'and C. T. Eddy, was presented by Mr. Eastwood. A. great many quantitative data were given, with little reference to the theory. Theoretical considerations were discussed by Dr. Gottschalk, and C. A. Edwards of the University College of Wales, contributed very interesting comment on the relationship of size and critical strain. Dr. Edwards discussed the necessity of considering strain with time and temperature.

Jan 1, 1935

By F. M. Jardine

I1944 the cry was for higher production more tons, more metal. New plants were built, capacity of old plants was increased and millmen all over the country were treating tonnages far above normal, sacrificing recoveries to vet that extra metal from extra tons. The list of "critical minerals," at first long, has been scratched until there remains, not a list, but only an item or two. Stocks of all metals and minerals, with possibly one or two exceptions, are now entirely adequate and numerous plants have curtailed or suspended operation.

Jan 1, 1945

By R. E. Allen

ONE of the most unusual oil field engineering accomplishments of the past two years is the development and rapid advance in the directed drilling of wells. Directed drilling as referred to herein is the controlled steering of the drill stem in the required direction to a given location in the oil sand. The deliberate drifting of holes is an act so far removed from what has been considered good drilling practice that all of the old questions of correlation, drillability and production difficulties are revived. Only a few years ago, during the boom days of Seminole and Santa Fe Springs, the crooked-hole problem, as it was then called, occupied a primary position among the topics of the day in the fields of California and the Mid-Continent. The net result of all the discussion and study at that time was the conclusion that oil wells should be drilled with as little deviation from

Jan 1, 1933

Edwin H. MessiteR, New Pork City (communication to the Secretary*):—Under the heading " Flues," Mr. Edwards refers to the Bee-hive construction, a cross-section of which is shown in Fig. 4 of his paper. A flue similar to this was designed by me about six years ago,' in which the walls though much thinner than those described by Mr. Edwards gave entire satisfaction. These walls, from 2.25 in. thick throughout in the smaller flues, to 3.25 in. in the larger, were built by plastering the cement mortar on expanded-metal lath without the use of any forms, or cribs, whatever, at a cost of labor generally less then $1 per sq. yd. of wall. Of course, where plasterers cannot be obtained on reasonable terms, the cement can be molded between wooden forms, though it is difficult to see how it can be done with an interior core only as stated by Mr. Edwards. In regard to the effect of sulphur dioxide and furnace-gases on the cement, I have found that, in certain cases, this is a matter which must be given very careful attention. Where there is sufficient heat to prevent the existence of condensed moisture inside of the flue, there is apparently no action whatever on the cement, but if the concrete is wet, it is rapidly affected by these gases. At points near the furnaces there is generally sufficient heat not only to prevent internal condensation of the aqueous vapor always present in the gases, but also to evaporate water from rain or snow falling on the outside of the flue. Further along, a point is reached where rain-water will percolate through minute cracks caused by expansion and contraction, and reach the interior, even though internal condensation does not occur there in dry weather. From this point to the end of the flue the roof must be coated on the outside with asphalt paint or other impervious material. In very long flues a point may be reached where moisture will condense on the inside of the malls in cold weather. From this point to

Jan 1, 1905

By Noel S. Christensen

COBALT is a silvery white metal with a slight bluish cast, strongly resembling nickel in its appearance and properties, notably its resistance to corrosion, although its alloys with other metals differ quite decidedly in their properties from those of nickel. The name cobalt was derived from the German "Kobold" meaning goblin and this came to be applied to the cobalt minerals which, although they resembled ores, did not readily yield metal when smelted by the primitive processes available in the Middle Ages. Cobalt minerals were used as pigments by potters and glass workers before the Christian era. Cobalt blue glass was found in the tombs of the Egyptians and in the ruins of Troy, and many colors used by the ancient Greeks and Romans contained cobalt. The ancients were entirely unaware that the colors they used were caused by cobalt and the metal itself was only isolated about the middle of the eighteenth century by Brandt although the preparation of cobalt compounds had been carried on in a small way ever since the opening of the silver-cobalt deposits at Schneeberg, Germany, in 1470. The metal is fairly hard, with a density when pure of 8.81 in the cast state and 8.92 after working. The commercial grade of cobalt has a density of about 8.75 in the cast state. The atomic weight is 58.97. The Brinell hardness of the pure metal is 124; that of the commercial grade varies from as low as 100 to as high as 138, depend¬ing upon the impurities present. Pure cobalt has a sharp melting point at 1478°C. The tensile strength of the pure metal in the cast condition is said to be higher than that of any other pure metal. Kalmus gives it as 34,400 lb. per sq. in. for cast unannealed metal; the effect of annealing at 850 to 900°C. is to raise the strength slightly. The yield point is close to the ultimate strength and the reduction in area and elongation are both low. Mechanical working raises the tensile strength of pure cobalt rapidly and in the form of wire it will be over 100,000 lb. per sq. in. The presence of small quantities of carbon in the commercial metal raises the tensile strength to as high as 61,000 lb. per sq. in., the reduction in area and elongation being raised at the same time to approximately 20 per cent.

Jan 1, 1933

By B. C. Mariacher

EXTRACTION of uranium from ores is being ac-complished by processes which. for the most part, subject the entire ore to acid or carbonate leaching. Ore deposits with a U 3 O 8 content below 0.10 pct U 3 O 8 are seldom considered suitable for treatment by leaching. A preliminary concentration that would enrich the uranium content of an ore by a simple, low cost process based on physical properties of the ore might result in some low grade deposits becoming commercial ores. In addition, the process might be employed in existing operations to reduce transportation and leaching costs and to increase capacity of existing leaching plants. A study to attempt the development of a preliminary concentration process for primary uranium ores was undertaken by the Colorado School of Mines Research Foundation under sponsorship of the U.S. Atomic Energy Commission. The objective of this work was to produce concentrates containing 0.25 pct U3O8 from the low grade ores tested. Ores Tested: The main effort was devoted to the low grade primary uranium ores from northwestern Saskatchewan. Samples were obtained from the Beaverlodge operation of the Eldorado Mining & Refining Ltd. Additional primary ores, obtained from deposits in Gilpin County, Colo., contained from 0.07 to 0.10 pct U3O8. Summary of Concentration Tests: The Beaverlodge ore was tested to determine amenability of the ore to concentration by magnetic, electrostatic, gravity, and scrubbing processes. None of these produced satisfactory results. Both gravity and magnetic processes produced fairly good concentrates when closely sized fractions of the ore were treated, but on the basis of treating the total ore, recovery was poor. Preparation of sized fractions and the low capacity of equipment for suitable concentration made these methods impractical. As flotation offered the advantage of treating the total ore without intermediate sizing, the main effort was in this direction. A flotation process was developed that fulfilled the concentration objectives as set by the AEC. Pilot plant testing was used to verify results obtained from laboratory batch testing. Mineralogy: A petrographic examination of the Beaverlodge ore included a study of polished sur- faces and identification of the radioactive mineral by autoradiograph and X-ray diffraction. Approximate quantitative mineral identification was as follows: quartz, 60 pct; orthoclase feldspar, 20 pct; chlorite, 10 pct; carbonates, 5 pct; and miscellaneous minerals, 5 pct. Included in this last group were plagioclase feldspar, pyrite, mica, chalcopyrite, pyroxene, sericite, magnetite, galena, and uraninite. The most general occurrence of uraninite was in the form of crusts and thin coatings on limonite-stained grains of pyrite, quartz, and pyrite-quartz intergrowth. At least 90 pct of the uraninite was still attached to other minerals in a 100 by 200-mesh size fraction. The uraninite crusts were as small as 10 to 20 µ diam, and 5 to 10 µ thick. The Flotation Process Petrographic examinations of the Beaverlodge ore had indicated the impracticability of attempting to concentrate the uranium by floating individual grains of uraninite. Liberation of the uraninite required grinding to sizes below those suitable for flotation. However, there was preferential association of the uraninite with some minerals while others were free of uraninite attachment. The approach to the development of a flotation process was, therefore, based upon an attempt to concentrate the uraninite by floating carrier minerals. The following paragraphs discuss the various stages of the process with regard to the factors tested and the conditions under which best results were obtained. Grinding: The most effective size range for flotation was —150 mesh + 13 µ. The —13 µ material in the final concentrate had a higher U3O8 content than the total ore, but not as high as the average concentrate; however, rejection of slimes before flotation was prohibitive because of the loss in uranium carried in the —13 µ fraction. Grinding techniques which contributed to a minimum production of fines, such as stage grinding, were then employed. Quartz and Silicate Depression: These minerals represented approximately 80 pct of the ore and were free to a large degree of uraninite attachment. Significant improvement in the grade of the concentrate was obtained by depression of these minerals with hydrofluoric acid or sodium fluoride. Promoter: Selective stage flotation of uraninite carrier minerals was simplified by development of a single promoter mixture. The mixture consisted of an emulsion of a fatty acid, fuel oil, and a petroleum sulfonate and was selected after a comprehensive series of tests. It contained three parts by weight of an oleic and linoleic acid such as Emersol 300,

Jan 1, 1957

By Yeatman, Pope

FOR many years geologists have felt that mining companies should adopt a more liberal policy in the publication of their reports. The increasing usefulness of the geologist to the mining profession in recent years demands that consideration be given to this question. Prior to 1900, the call for geologists of high reputation by mining companies was probably mainly in connection with lawsuits arising from the interpretation of rights or ownership as defined by the "Law of the Apex," and from problems such as finding the continuation of the vein or whether the vein followed was the main one or a branch. Later, outside of legal questions, the most intensive and general use of their services arose from the search for oil. Geologists to determine the position of the formations

Jan 1, 1938

By R. J. Reynik

A semiempirial small flunctation theory of diff- sion in liquids is presented, which employs a fluctuation energy assumed quadratic for a small atomic or molecular displacement and Einstein's random-iralh model. The resulting diffusion equation is given by In these equations. D is the diffusivity, is the average liquid shite coordination number (at interatomic distance d. cm. T is the absolute temperature, xu. em, is (the diffusive displacement. K, is the quadratic fluctuation energy force constant, and rg, cm, are the radii oj diffusing atoms A and B, respectively. The quantities Xn and K are calculated from the computer-filled values of the slope and intercept. respectively. The radius of self-diffusing atom or radii and of diffusing atoms A and B are eta United and compared with values reported in the literature.. The predicted linear variation of diffusivity with. It tempera lure htm been observed in approximately thirty-iire metallic liquid systems, and in over seventy-fiee other liquid systems, including the organic .alcohols, liquified inert gases, and the molten salts, ALTHOUGH the average density within a macroscopic volume element of liquid is constant for fixed total number of atoms. pressure. and temperature, there exist microscopic: density fluctuations within the respective volume element. As such the microscopic volume available to an atom and its Z first nearest neighbors at any instant of time fluctuates above and below the average volume available to these atoms. If one assumes that liquid state atoms vibrate as in a solid. and further postulates that the mean position of any atom in the liquid state is not stationary. but shifts during every .vibration a distance 0 5 j 5 xo. then every atom in the liquid state continuously undergoes diffusive displacements which vary in the range 0 5 j 5 ro. Mathematically. for a binary liquid system consisting of atcrms A and B. the maximum diffusive displacement. .YO, is defined by the equation: where d is the average liquid state interatomic distance at specified liquid state coordination number Z. and v~ \ and vg are the effective radii of diffusing atoms A and B: respectively. For self-diffusion. r^ equals rg , and Eq. [I.] reduces to: It is interesting to note that Eq. [l] or [2] can be used to compute the radii of the diffusing atoms, provided one had an experimental evaluation of xo. As such. the computed radii could be compared with metallic or crystallographic ionic radii to ascerlain the electronic character of the diffusing atoms. Thus it is proposed that in the liquid state the n~otion of an atom relative to its original equilibrium position of oscillation represents the thermal vibration of any atom and its Z first nearest neighbors. while the small and variable displacements. 0 5 1 5 xc,. of the centers of oscillation represent the complex diffusive motions of the atoms at constant temperature and pressure. This is consistent with data obtained from slow neutron scattering by liquids1 ' and resembles an itinerant oscillator model of the liquid state.'" It is further postulated that the atomic displacements characterizing the liquid state diffusion process are essentially a random-walk process. As such. it nlay be described by Einstein's equation:' where D is the diffusivity. sq cm sec-'. j2 is the mean square value of the diffusive displacement. and i> is the frequency of density fluctuations giving rise to diffusion. FORMULATION OF DIFFUSION EQUATION The effective spherical volume occupied by an atom, as a consequence of a microscopic density fluctuation which enlarges the volume available to any atom, exceeds its average liquid state atomic volume by an amount: where AV is the enlarged spherical volume, v is the radius of the diffusing atom. and j is the elementary displacement distance from the original center of oscillation of the vibrating atom to a new center of oscillation position. For small atomic displacements. where c is a constant whose value depends upon the assumed geometry of the enlarged volume. For a spherical increase in volume, c equals 4nr2. Following the treatment of Furthl' and ~walin." assuming the enlarged volu~nes AL7 for the diffusing atoms are distributed in a continuunl. the probability of finding a fluctuation in the size range 0 5 j 5 xo defined by Where c includes the geometric constant cl and Eij) is the fluctuation energy causing the volume change. But the proposed model assumes all the Z first nearest-neighbor atoms are centers of oscillation. and hence the probability that any of these atoms is adjacent to a fluctuation of magnitude 05j5xo is unity. Thus:

Jan 1, 1970

By R. E. Ogilvie, N. L. Peterson

Diffi-lsion measurements were conducted at all compositims in the bcc solid solution of the U-Nb system employing incremental couples at composition intemals of 10 at. pct. Diffusion coefficients were determined by the Matano method from concentration gradients obtained with the electron-probe microanalyzer. The activation energy for inter-diffi-lsion as a function of compositim shows three distinct regions: 1) 80 to 100 pct U.6= 30 kcal per mole; 2) 20 to 80 pct U, $ - 70 kcal per mole; 3) Oto 20 pet U, Q = i40 kcal per mole. The frequency factor, fi0 and the activation energy $ were found to be roughly related by the following equation: log Do ^9.7 X IO-5Q -6,6. The Kirkendall marker movement indicates that DU is larger than DNb between 16 and 100 pct U and DNb is larger than DU from 0 to 4 pct U. FOR practical as well as fundamental reasons, the rates of diffusion in alloys are of considerable consequence. Most solid-state reactions are largely dependent upon the diffusion of atoms through the lattice structure and along grain boundaries. The high-temperature strength and reasonable nuclear properties of niobium have prompted its use as a reactor material in contact with uranium fuel. Hence, diffusion data for the U-Nb system are of considerable importance. In the previous diffusion study1 on the U-Nb system using pure element couples, reliable data were obtained only in the range of 0 to 10 at. pct Nb due to the large variance of the diffusion coefficient with composition. Also, a large Kirkendall effect and considerable porosity in the uranium-rich areas of the specimen were reported, which suggests that the true diffusion coefficients are somewhat larger. The purpose of the present study was to obtain reliable diffusion coefficients at all compositions using incremental diffusion couples with intervals of 10 at. pct. In view of the abnormal self-diffusion be- havior of y uranium2-4 and some other bcc transition elements,&apos;-&apos; it was felt that a comparison of the interdiffusion coefficients in the bcc U-Nb system with those of Reynolds et al.9 for the fcc gold-nickel system might shed some light on the diffusion mechanism involved. Both systems have similar phase diagrams, in that complete solid solubility exists above a miscibility gap. EXPERIMENTAL PROCEDURE The uranium used in this investigation was obtained through the courtesy of Argonne National Laboratory. An analysis of this material detected only Si-30, A1-7, C-6, N < 10 and 0-18 ppm. The niobium was electron-beam melted material obtained from Stauffer-Temescal. The gaseous impurities were less than 50 ppm, and the spec troc hemical analysis showed Ta-500 and W-200 ppm. U-Nb alloys were prepared at composition intervals of 10 at. pct by melting the appropriate amounts of the pure elements in an arc furnace. The buttons were inverted and remelted 6 times to assure complete mixing. The buttons were then wrapped in molybdenum foil, canned in Zircaloy-2 or stainless steel, and hot rolled 30 pct reduction in thickness at temperatures between 850" and 1100°C. Alloys with 10, 20, 30, 40, and 90 at. pct Nb rolled quite easily under these conditions, but the 50, 60, 70, and 80 pct alloys remained brittle. After melting and rolling (when possible), the alloys were annealed for 24 hr at a temperature within 100°C of their melting point in a dynamic vacuum of better than 4 x 10-8 mm Hg. These treatments produced alloys which were homogeneous on a 1 p scale within the detectability limits of the electron probe. During fabrication, the alloys picked up as much as 100 ppm Mo and 100 ppm Zr. Other elements checked for but not found were Co, Cr, Fe, Mn, Ni, and Ti. The grain size of the annealed samples ranged from 3 mm for the uranium-rich alloys to 0.3 mm for the niobium-rich alloys. This permitted measurements of the concentration gradients in the diffusion samples without crossing more than one or two grains, thereby eliminating any grain boundary effects. The specimens were bonded by theU&apos;picture frame" technique as reported by Kittel.10 Specimens of composition b)U + (100 - x)Nb were sandwiched between two specimens of composition (x + 10)U + (90 - x)Nb after they were ground flat and parallel

Jan 1, 1963

By J. Chipman, N. J. Grant, B. M. Shields

The vacuum tin-fusion method of analysis for hydrogen, developed by Carney, Chipman, and Grant, has been modified to permit the analysis of the evolved gases for hydrogen by means of a thermal conductivity cell. A properly prepared sample can be analyzed in 10 min with a probable error of ±0.12 ppm. A study of various methods for storage of hydrogen samples shows that samples can be safely held in a dry ice-acetone bath as long as six days. Storage in liquid nitrogen is necessary for samples to be held one week or more. HE vacuum tin-fusion method, as developed by I- Carney, Chipman and Grant,&apos; is the only analytical procedure which has shown promise of being fast enough for use in the control of hydrogen during steelmaking. It was felt that further simplification and faster speed of operation could be effected by the use of thermal conductivity measurements for analysis of the gases evolved in the tin-fusion method. The application of conductivity measurements to the tin-fusion method is possible because: 1—the evolved gas is essentially a mixture of hydrogen, nitrogen and carbon monoxide with a hydrogen content usually over 50 pct, 2—the evolved gas is collected at a relatively low pressure, and 3— the thermal conductivities of CO and N2 are practically identical while that of hydrogen is very much greater. The major part of this research program was devoted to the construction and calibration of a vacuum tin-fusion apparatus which analyzes the evolved gases for hydrogen by means of a thermal conductivity cell. The second phase of the problem was associated with the development of a procedure for storage of samples prior to analysis. With the rapid quenching method for hydrogen sampling,&apos; which seems to be the most practical for steel mill use, it is necessary that the samples be stored safely during the interval between sampling and analysis if the hydrogen content of the molten metal is to be maintained in the supersaturated solid samples. The thermal conductivity bridge has been used for a number of years in the analysis of certain gas mixtures. An elementary discussion of the theory and practice of gas analysis by thermal conductivity measurements is given by Minter.3 A more comprehensive discussion of the theory and of the various measuring circuits is presented by Daynes.&apos; A complete knowledge of the theory and properties of the thermal conductivity of gases and gaseous mixtures can be gained by a study of the standard textbooks on the kinetic theory of gases."&apos; The existing data on the thermal conductivity of single gases are reviewed by Hawkins: that for a number of binary gas mixtures by Daynes&apos; and Lindsay." The thermal conductivity method may be applied to the determination of the composition of a binary mixture if: 1—the thermal conductivity of the mixture varies monotonically with composition, and 2— the two gases have measurably different thermal conductivities. The greater the difference between the two gases, the greater the sensitivity of the method.10 he method is applicable to the analysis of multicomponent mixtures when all of the gases in the mixture except one have nearly the same thermal conductivity. Fortunately, the mixture of hydrogen, nitrogen, and carbon monoxide evolved by the tin-fusion analysis&apos; falls in this latter classification. The thermal conductivities of nitrogen and carbon monoxide are practically equal; and the thermal conductivity of hydrogen is approximately seven times that of the other two. Therefore, the thermal conductivity of a gaseous mixture of hydrogen, nitrogen, and carbon monoxide at known temperature and pressure can be related directly to the percentage of hydrogen in the mixture by suitable calibration. Usually the thermal conductivity of a mixture of gases is measured at atmospheric pressure where the thermal conductivity is independent of pressure over a wide pressure range. At very low pressures (below 1 mm Hg), the thermal conductivity of gases varies with the pressure. This phenomenon has been utilized in the Pirani vacuum gage for the measurement of pressures in the range of 10" to 10-0 mm of mercury.= Very little has been published concerning the variation of thermal conductivity with pressure at intermediate pressures between 1 mm Hg and 1 atm. However, preliminary measurements indicated that the thermal conductivities did vary with pressure over the range of pressures (up to 10 mm Hg) at which gases are delivered from the vacuum pump. Therefore, the calibration of the thermal conductivity cell had to be planned to include the effects of both gas composition and pressure. Such a calibration chart is shown in Fig. 4. Most industrial applications of the thermal conductivity method of gas analysis have used a compensated Wheatstone bridge circuit containing two

Jan 1, 1954

By F. R. Jones

AT Tahawus, N. Y., National Lead Co. operates the MacIntyre development. Here the world&apos;s largest titanium mine produces 5200 long tons of ore per day and pours 8000 long tons of waste rock over its dumps. Concentrated ilmenite is sent by rail to National Lead Co. pigment plants, and a second product, magnetite, is sold to steel producers in raw form or is agglomerated and shipped as sinter. Several earlier attempts had been made to produce iron from the deposits, which have been known since 1826. These attempts failed, chiefly because of titanium impurity. In 1941 the present owners reestablished the operation for production of war-scarce ilmenite, and the impurity became the main product. The Ore: The MacIntyre ore zone is about 2400 ft long and 800 ft wide in horizontal measurements. Ore outcrops were found on the northwest side of Sanford Hill, 450 ft above Sanford Lake and 2500 ft southeast. The zone dips at about 45" toward the lake and plunges to the southwest. The ore minerals, ilmenite and magnetite, are unevenly distributed in bands roughly parallel to the long axis of the ore zone and are interspersed with bands and horses of waste. Hanging wall ores are fine grained and grade from rich ore to waste rock or gabbro. Footwall ores are coarse grained and are almost entirely ilmenite and magnetite. The foot-wall waste rock, anorthosite, is the common country rock. Several faults cut the ore zone. These faults have no great displacement but do contribute to the great physical variations in ore rock and surrounding waste. The Mine: The MacIntyre mine is an open pit operation, with benches at 35-ft intervals. The lowest bench is now 54 ft below lake level. Loading equipment consists of three electric-powered shovels (a P & H model 1400 with 4-yd dipper and two Bucyrus-Erie models 85-B with 2%-yd dippers) and one diesel-powered shovel (a Northwest model 80D with 2%-yd dipper). Ore and waste are transported to a 48x60-in. jaw crusher in ten 22-ton Euclid trucks with 300-hp diesel engines. Ordinarily the two Bucyrus-Erie 2 % -yd shovels load ore into a fleet of three or four trucks. This combination works two 8-hr shifts per day, moving 5200 long tons of ore to the crusher and removing a small portion of the waste rock. The P & H model 1400 shovel, with a fleet of four trucks, loads waste on three shifts per day. The mine operates on a 5-day week, with a small maintenance crew working Saturday. Oversize rock is broken by a dropball handled by an Osgood model 825 rubber-mounted crane.&apos; Ore and waste are broken by drilling and blasting 9-in. diam vertical holes behind the benches. Bucyrus-Erie 42-T churn drills are used to drill the holes, which are extended 4 ft below the bench level on which the broken rock will fall. Drilling and Blasting History: In its early years the mine was equipped with Bucyrus-Erie 29-T churn drills, which drilled 6-in. holes. To keep up with production requirements the hole diameter was soon increased to 9 in., and by 1950 the three 42-T drills now in use had been acquired. Early blasting experiments with different kinds and grades of explosive led to adoption of 90 pct straight gelatin dynamite as standard. It was recognized that this explosive was expensive, and from the start of operations until 1950 extensive experiments were made using blasting agents of the ammonium nitrate family. Results were recorded as uniformly poor, with great build-up of oversize rock. The expense of these experiments, and the discouraging results, caused the abandonment of any expectation of breaking MacIntyre rock with anything but 90 pct straight gelatin dynamite. Further standardization led to 9-in. well drillhole spacings set at 16 ft in ore and 18 ft in waste, exceptions being permitted only for unusual conditions. The hole burdens were theoretically about 22 ft. Due to the extreme back-slope of bench faces, caused by blasting with heavy charges of dynamite, actual burdens were commonly well over 30 ft. Lack of precise control resulted in many holes having a burden as light as 15 ft. General practice was to stem 6 or 7 ft of hole with magnetite concentrate, the amount of stemming being left to the discretion of the pit foreman. Usually all holes in a row were fired instantaneously with Primacord detonating fuse. Millisecond delays were

Jan 1, 1957

By Werner Schwartz, Wolfgang Haase

Lead sulfide concentrates, which may include other lead concentrates, are sintered on an up-draught sintering machine without the addition of any diluting agents or fluxes. Subsequently they are melted in an oil- or gas-fired rotary furnace. The sintering and melting processes are based upon the following roast-reaction: PbS + 2 PbO = 3 Pb + SO, PbS + PbSO, =2 Pb + 2 SO, For obtaining a lead bullion free from sulfur, the sintering process is carried out in such a way that the sinter product contains a small amount of excess oxygen above that to react with the sulfides. At the end of the melting process, when the reactions are finished, the remaining small amount of oxide residues is reduced with coal to which a certain percentage of soda ash (about 1 pct of the lead bullion) is added. For the lead smelting process described neither coke nor fluxes—except soda ash—are required. This process is being utilized by a European smelter successfully and with a high lead recovery. The consumption figures for the smelting of 100 tons per day of lead concentrates are indicated. The lead content of the lead concentrates from modern ore dressing plants ranges from 65 pct to above 80 pct. In most lead smelters of the world these concentrates are smelted in a blast furnace. For blast-furnace smelting the concentrates have to be desulfurized and agglomerated by sintering. A requirement for the perfect operation of a down-draught sintering machine and of a blast furnace is a maximum lead content in the feed of 40 to 45 pct. For this reason, some lead concentrates have to be diluted by adding return slags, limestone, and possibly iron oxide and sand. As an example, 100 tons of lead concentrate with 72 pct Pb would contain 13.5 tons of gangue (including the zinc). To produce a perfect sinter with 42 pct Pb it would be necessary to add 70 tons of flux and return slag, more than five times the original weight of the gangue, to the sinter mix and blast-furnace charge. A correspondingly large amount of coke would be required in order that all of these materials reach the heat of formation and the melting temperatures of the slag (1200" to 1400°C) inside the blast furnace. The roast-reaction process presents a possibility for lead recovery without dilution of the concentrates. In this process the concentrate mixed with coal is placed upon a Newnam-hearth and air is blown through nozzles into the heated mix. AS a result metalllic lead and a relatively great amount of so-called .&apos;Grey Slag" with a lead content of 25 to 35 pct are formed. The slag is sintered to eliminate sulfur and, after addition of the requisite fluxes, treatt:d in a blast furnace. Owing to the poor recovery of lead from the hearths and to the unavoidable heavy hand-work plus the risk of poisoning this process is utilized in very few 112ad smelters today. Since in mxny countries of the world coke is expensive and difficult to obtain, it appeared feasible to use the principle of the roast-reaction by modern sintering and melting methods with recovery of the lead in electric, or oil, gas, or coal-fired furnaces. Two processes are utilized on an industrial scale: A) Lead smelting in the electric furnace of the Bolidens Gruv A/B in Sweden, as described by S. J. Walldcn, N. E. Lindvall, K.G. Gorling, and S. Lundquist. B) The self-fluxing lead smelting of Lurgi Gesell-schaft fiir Chemie und Huttenwesen m.b. H., Frankfurt a M, Germany, which is described in this paper. In the Boliden process referred to above the sinter mix is pelletized by enveloping return fines with layers of flue dust, limestone powder, and dried galena concentrate. The roasting and agglomeration are carried out on a down-draught machine, and a slight excess of sulfur is left in the sinter product. During the smelting in the electric furnance the roast-reactions occur and a slag poor in lead and a sulfur bearing lead are formed. This latter is subsequently oxidized in a converter to obtain lead bullion and dross. The Lurgi-process achieves the maximum possible extent of the roasting reaction on the sintering machine. The wet flotation concentrates are blended with return fines (lead content 70 to 80 pet), any existing flue dusts and lead slimes—but without the

Jan 1, 1962

By F. H. Buttner, H. Udin, J. Wulff

Using a modified Udin, Shaler, and Wulff technique, the surface tension of gold Udin, purified helium was found to be 1400 ± 65 dynes per cm for the temperature range 1017° to 1042°C. IN the original Udin, Shaler, and Wulff technique for measuring the surface tension of copper: variously weighted wires were allowed to extend or contract in a copper cell held at elevated temperatures in vacuum. By plotting stress vs. strain for a wire array in one test, the stress at zero strain is obtained. This is the point where the contractile forces resulting from surface tension are balanced by the applied load, according to the expression: y = e=o r [1] where y is the surface tension in dynes per cm; a,,,, the stress at zero strain in dynes per cm; and T, the radius of the wire in cm. The assumption that the wires deform viscously permits the drawing of a straight line through the points on the stress-strain plot. Justification of the assumption has received further experimental support recently.&apos;-&apos; The presence of grain boundaries in the wires requires a correction to the original expression used." Thus: y = d4=T [l- (dl) (ar)Y1 [2] where, n/l is the number of grain boundaries per unit length, and a, the ratio of grain boundary tension to free surface tension. Alexander, Kuczynski, and Dawson in studying the creep of gold wire in vacuum were unable to obtain reproducible values of the surface tension of gold. In plotting stress vs. strain for progressively longer times, they found that the stress at zero strain drifted with time from positive stress values to negative values. Similarly, for the surface tension of silver, reproducible values were obtained only when a purified helium atmosphere was substituted.&apos; Evidently the evaporation rate of silver in vacuum is too high at the temperatures employed to obtain solid-gas equilibrium even in a similar metal enclosure. Thus reproducibility of results is lost. Experimental Procedure The experimental procedure was much the same as that originally developed by Udin, Shaler, and Wulff with a few modifications and improvements. For greater accuracy in strain measurements, knots gave way to cut gage marks as shown in Fig. 1. These were made with a hand-driven lathe in which razor blades serv&apos;ed as cutting tools. Also a more precise cathetometer with a screw accurate to 0.00015 cm was used. The tests were conducted in an atmosphere of purified .helium rather than in vacuum in order to avoid possible evaporation difficulties. Five mil wire of high purity gold (99.98 pct) was used. After cutting in the gage marks, each wire of a series of about 12 was differently loaded by welding a gold ball to one end. This was done by dipping the end of the wire in a cooling gold droplet, previously melted on a charcoal block with a No. 2 acetylene torch. The other end of the wire was strung through a hole in a gold lid and twisted over the edge to hold the wires fixed and in suspension from the lid. The lid and mounted wires were then dipped in pure ethyl alcohol to dissolve any skin oils and dirt on the surface of the wires due to handling. Finally the lid was put in place on an alundum crucible lined with gold so that the wires hung freely within the gold-lined chamber. This whole assembly was next heated in a quartz nichrome wound tube furnace and heated for a few minutes at 600°C to soften the wires. After this anneal the wires were easily straightened with tweezers. The wire assembly was finally annealed 10" to 25°C above the subsequent test temperature for 2 hr. This treatment allowed the grains to grow to equilibrium size and shape. After the anneal, the lid was mounted in front of the cathetometer. The gage length was measured by sighting the 40 power microscope on the upper lip of the lower gage mark for the first reading, then traveling up to the lower lip of the upper gage mark for the final reading. This procedure was repeated four times to give an average gage length value. In this manner the annealed gage length and the final gage length could be measured to determine the strains. During all measurements, grain counts were made.

Jan 1, 1952