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By R. W. Heckel, R. L. DeGregorio

The DeHoff method of determining the size distribution of ellipsoid-shaped, second-phase particles has been applied to the spheroidization of cementite in a eutectoid steel. The surface area of Precipitates determined from the various size fractions in the distribution was correlated with surface-area measurements based on the number of intersected interfaces on a random lest line. The precipitate particles were found to be oblate ellipsoids with an axial ratio of about 0.90. The size distributions were found to be log-normal. A method is proposed whereby the shrinkage of small particles and growth of large ones can be determined from the experimental data. The experimental data are compared to previously proposed mathematical models describing diffusion-controlled kinetics and various types of interface-controlled kinetics. The experimental growth and shrinkage rates are considerably slower than those predicted by diffusion-controlled kinetics. The best fit is obtained for a model describing interface-controlled kinetics limited by the rate of formation of cementite at the growing interfaces where the interfacial reaction rate is proportional to the solute thermodynamic activity gradient across the surface. THE subject of spheroidization of second-phase particles has been considered previously by other investigators. Livingston1 has studied the precipitation of the cobalt-rich phase from copper alloys containing 0.7 to 3.2 wt pet Co. His results indicate that the average particle size increases as the cube root of the heat treatment time in accord with diffusion-controlled kinetics.'-' Komatsu and rant' in their studies of the growth of SiO2 in a dispersion-strengthened copper-silica alloy found that the initial growth of SiO2 proceeded by diffusion-controlled growth and later stages were limited by interface-controlled growth. The values of the activation energy obtained for the interface- controlled process led them to the conclusion that the process was limited by the dissociation of SiO2 at the shrinking interface. The transition from diffusion-controlled growth to interface-controlled growth was characterized by a particle size which varied with the heat-treatment temperature. It is also interesting to note that the particle-size distribution as found by Komatsu and Grant exhibits log-normal behavior when replotted on log-probability coordinates. Dromsky, Lenel, and Ansell9 have observed the growth of Al2O3 particles having a mean free path of about 1.5 to 13 µ. Their photomicrographs indicate that the Al2O3 particle sizes were larger than those observed by Komatsu and Grant. Dromsky, Lenel, and Ansell concluded that the Al2O3 coarsened by an interface-controlled growth mechanism limited by the solution of A12O3 at the shrinking interfaces. Bannyh, Modin, and Modin10 have studied the spheroidization of a eutectoid steel (0.83 wt pet C). Their spheroidization (tempering) treatments were carried out in the range from 210o to 700°C for times between 1.5 sec and 20 hr. Measurements of mean particle size as a function of time indicated a cube root of time dependence of size at 700°C, in agreement with previous analyses of diffusion-controlled kinetics.2°7 At lower temperatures, the time dependence was less than the one-third power. It is important to note that, although mathematical models of growth* considering particle-size distribution have been available, measurements of only mean particle size have been carried out. DeHoff11 has presented a quantitative metallography technique which is applicable to the determination of the size distribution of ellipsoids of constant shape. This method is applicable to both oblate and prolate ellipsoids of all ratios of minor to major axis and is based upon an extension of Saltykov's analysis12 of the distribution of spheres of varying size. DeHoff's analysis is based upon measurements of the minor axis of the particles on a random plane of polish in a unit area. This method provides a measure of the number of ellipsoids in various size ranges per unit area. As pointed out by DeHoff, such measurements may be used to obtain surface area and total precipitate volume data. Thus, the accuracy of the distribution analysis may be checked by comparison of the surface and volume

Jan 1, 1965

By AIME AIME

WHEN "Chief" Buehler in mid-September set out on his official 10,000 mile swing-around-the-circle visiting Local Sections he decided not to tell his audiences how to organize and operate a state geological survey. As the director for 28 years, of the Missouri Bureau of Geology and Mines (which, he concedes, is the best in existence) lie could have done this to good advantage; but he preferred to discuss affairs of the Institute. In 25 speeches made to member and student groups, he dwelt on sonic or all of the following phases of activities. He stressed the vital importance of the Local Sections in providing the opportunity for members to associate under pleasant auspices with others in their profession. The benefits of Institute membership he classified as tangible and intangible. Of the two lie felt that the latter were more important. A man's progress in his profession depends in large measure upon a broad acquaintance and friendship with others. The Institute is incomparably the best medium for meeting fellow engineers. The primary, function of the new enlarged Membership Committee is to obtain a comprehensive list of engineers eligible to be members, always hearing in mind that the standards of membership must not be lowered in any degree. These men are to be invited by individual letter to apply for membership. Whether they become members or not is of less importance; but it is vital that no qualified man should be outside the fold for the reason that membership had never been suggested to him. Dr. Buehler emphasized repeatedly his feeling that the younger man that had never had Institute membership brought specifically to his attention had been cheated:

Jan 1, 1935

By J. D. Price, W. M. Bertholf

In preparing a paper on coal washing in Colorado and New Mexico, it is difficult to refrain from entering into a discussion of the historical aspects of this subject, for the story of coal washing in these states goes back well into the last century. This phase of the subject could well make a complete paper in itself, and of necessity this present discussion must be limited to current operations. Therefore, we are taking into consideration only those washing plants which have operated for some period during the years of 1947 and 1948.

Jan 1, 1949

No. Place Date Vol. 1. Wilkes-Barre, Pa May, '71 1 2. Bethlehem. Pa Aug., '71 1 3. Troy, N. Y Nov.,'71 1 4. Philadelphia, Pa Feb., '72 5. New York, N. Y.* May, '72 8. Pittsburgh, Pa Oct., '72 7. Boston, Mass Feb., '73 8. Philadelphia, Pa.* May, '73 9. Easton, Pa Oct.. '73 10. New York, N. Y.Feb., '74 11. St. Louis, Mo.* May, '74 12. Hazleton. Pa Oct., '74 13. New Haven, Conn Feb., '75 14. Dover, N. J.* May, '75 15. Cleveland, O Oct., '75

Jan 1, 1929

By R. T. Hukki

The purpose of this paper is to present an analysis of the unit operation of grinding and the circulating load, of the unit operation of classification and the circulating load, and of the two superimposed into one operation to clarify the overall trend of events relating the circulating load to the capacity of the closed circuit grinding system, to the specific surface area of the final fine product, and to the energy consumption in the closed circuit grinding process. The capacity of the closed circuit is frequently, if not always, determined by the classifier, not by the mill." With these words a startling idea was expressed in 1934 by Dorr and Anable.1 Today, some 30 years later, grinding in closed circuit is universally accepted. The basic curve presenting the capacity factor versus circulating load by Dorr and Anable has obviously been so convincing that not a single authoritative investigation regarding its validity seems to have been conducted. The only major subject of discussion appears to have been the relative merits of the 'high' and 'low' circulating loads. It seems to be universally agreed that much still remains to be hoped for regarding sharpness of classification. BASIC CURVES FOR CONSTANT MILL LOADING Fig. 1 presents a series of basic curves for constant mill loadings in function of new feed rate and circulating load. The mill loading indicated by each curve is the sum of hourly tonnages of new feed and circulating load. For example, at point 0 where the new feed rate is 100 t/h and the circulating load 300%, or 300 t/h, the mill loading is 400 t/h. In closed circuit grinding, the new feed rate corresponds to the rate of finished classified material produced. Although at this stage the effect of classification is excluded in construction of the basic curves, the trend indicated by these curves is strongly in favor of closed circuit operation under relatively small circulating loads. A study of the curves reveals that, e.g., a mill operating under a load of 200 t/h at point P where the new feed rate is 50 t/h and the circulating load 300% will also operate under the same total load at point Q where the new feed rate is 100 t/h and the circulating load 100%. In other words, a reduction of circulating load from 300% to 100% would lead to doubling of the new feed rate and doubling of the rate of production of finished material, without a change in the total mill loading. The obvious question is, is it also possible by a reduction of the circulating load in a closed industrial grinding circuit to increase the production of finished classified material and, if so, how and to what extent? GRINDING TESTS BY DORR AND ANABLE In a series of open circuit grinding tests reported by Dorr and Anable1 limestone was ground at various feed rates in a 3-ft ball mill. The original data are reproduced in Table I. It can be seen that with the increased feed rate production of -65-mesh material increased and the energy consumption per unit weight of the -65-mesh product decreased in the reverse ratio. The data shown in Table I have been reproduced for our purposes in Table II for a scale corresponding to grinding in a large industrial mill. This reproduction

Jan 1, 1968

By John Lohrenz

Consider an investor who, knowing future costs and revenues, can choose how rapidly to produce from a given mineral source. If the investor elects to make that choice to maximize present value of future profits, how do policies affect the choice? Various alternative policies are examined. Some, such as higher royalty rates, lead to slower production rates than would otherwise obtain. Clearly, the chosen rates of production are policy-dependent. One is wisely counseled to consider how mineral policies affect the rate of production from given sources when analyzing current and alternative policies.

Jan 1, 1982

By Earle C. Waite

THE Industrial Minerals Division this year held no individual sessions of its own except the business meeting Tuesday afternoon. One session was held jointly with the Society of Economic Geologists, one with the Milling Methods Committee, and two with the Society of Economic Geologists and the Mining Geology Committee. At all of these meetings the attendance was better than usual and the Tuesday afternoon session, with nearly 400 present, probably set a new record for the Division. Joint sessions seem extremely popular when papers of mutual interest are presented. On Monday afternoon, mica was the predominating subject. The North Carolina mines provided most of the information presented, although the discussion brought out similarities with the New England mines. B. C. Burgess particularly stressed the fact that with the existing curtailment of mica from India it is necessary to get the most mica in the shortest time from the mines in this country. He said that the Colonial Mica Corp. hopes to have 1000 mines operating in the near future. This seems a great number, but when one considers that these mines employ from 3 to 28 men, mostly nearer the low figure, the output of each mine is relatively small. Although the price of mica is from six to ten times the prewar price, the yield of 8 to 150 lb. of mica per ton of raw ore does not provide high returns to the producer.

Jan 1, 1943

By AIME AIME

THE petroleum engineers have the conference habit. They drop in, thresh things over, and drop out. No time is wasted. So it was at the Fort Worth meeting of the Petroleum Division, Thursday and Friday, Oct. 8 and 9; headquarters, Blackstone Hotel. Associate Chairman Albertson arrived the morning of the 7th, to tie in any loose ends of the technical program. S. M. Greenidge, chairman of the local committee, checked the arrangements with his committeemen at a luncheon that noon. By mid-afternoon the next day's attendance was anybody's guess-there were no noses to count. But when the gavel sounded Thursday morning 175 members and guests were present, also the official A.I.M.E. party consisting of President Lovejoy, Secretary Parsons, and George Otis Smith. Shortly after adjournment Friday afternoon all but a handful had vamoosed, lit out, disap¬peared. But, in the meantime, it was a grand conference, attendance maximum every session, and the discussion fast and furious. The discussion came from all over the floor, sometimes as good-natured sniping, sometimes in long volleys that were almost fatal to the reporter. Everyone knew his subject and few there were who had nothing to say.

Jan 1, 1936

By Thomas Henderson

Strategically, intermaterials competition is important to companies involved in raw materials extraction, smelting, refining or other processing, semi-finished and finished goods fabrication, and end-use consumers. In other words, just about every company involved in manufacturing. The array of materials covered by other papers in this session is testimony to the subject's breadth. In trying to provide an over- view of such a broad topic, the purpose has been to relate some of the common characteristics of the markets, economics and technology of raw materials, and to describe some of the ways in which Intermaterials competition impacts the markets, economics and technology of raw materials.

Jan 1, 1982

By J. B. Dyess, H. A. Miley

Tarnish films on some of the common metals (particularly on copper and silver) have been of much scientific and commercial concern for a long time, but before the development of the electrical method1 of measuring film thicknesses there was no satisfactory quantitative approach by which to study the separate parts of these complex films. This method has been adapted to the measurement of duplex (cuprous-cupric) oxide films on copper.2 In the present work it has been used to measure tarnish films on copper when they were mixtures of cuprous oxide and cuprous sulphide, and also when they contained cuprous oxide, cupric oxide and cuprous sulphide, each component part being measured at a different potential value. Oxide-coated copper specimens, prepared by heating copper in air at chosen temperatures for given periods, were exposed to moist hydrogen sulphide in air to forml the various oxide-sulphide films studied. The comparative resistances afforded by (I) different thicknesses of oxide films, and also (2) equal thicknesses of oxide and sulphide films, against severe attacks of hydrogen sulphide were measured. The electrical method requires that the time be observed for the cathodic reduction of each part of the film substance, and these respective time values multiplied by the constant current give the corresponding quantities of electricity expended. Then each millicoulomb per square centimeter represents a film thickness of M X 108____ 2DX 96,494 X 103 where M is weight of oxide (or sulphide) giving one gram-atom of oxygen (or sulphur) for cathodic reduction, D is density of the oxide (or sulphide) and t is the thickness in Angstrom units per millicoulomb per square centimeter. The value of t for cuprous sulphide (cubical) is 14.27, and is obtained by substituting 5.78 for D and 159.20 for M in equation a. Previous values hf t for cuprous and cupric oxides (12.36 and 6.39, respectively) were used in the present work.

Jan 1, 1939

By J. B. Dyess, H. A. Miley

Tarnish films on some of the common metals (particularly on copper and silver) have been of much scientific and commercial concern for a long time, but before the development of the electrical method1 of measuring film thicknesses there was no satisfactory quantitative approach by which to study the separate parts of these complex films. This method has been adapted to the measurement of duplex (cuprous-cupric) oxide films on copper.2 In the present work it has been used to measure tarnish films on copper when they were mixtures of cuprous oxide and cuprous sulphide, and also when they contained cuprous oxide, cupric oxide and cuprous sulphide, each component part being measured at a different potential value. Oxide-coated copper specimens, prepared by heating copper in air at chosen temperatures for given periods, were exposed to moist hydrogen sulphide in air to forml the various oxide-sulphide films studied. The comparative resistances afforded by (I) different thicknesses of oxide films, and also (2) equal thicknesses of oxide and sulphide films, against severe attacks of hydrogen sulphide were measured. The electrical method requires that the time be observed for the cathodic reduction of each part of the film substance, and these respective time values multiplied by the constant current give the corresponding quantities of electricity expended. Then each millicoulomb per square centimeter represents a film thickness of M X 108____ 2DX 96,494 X 103 where M is weight of oxide (or sulphide) giving one gram-atom of oxygen (or sulphur) for cathodic reduction, D is density of the oxide (or sulphide) and t is the thickness in Angstrom units per millicoulomb per square centimeter. The value of t for cuprous sulphide (cubical) is 14.27, and is obtained by substituting 5.78 for D and 159.20 for M in equation a. Previous values hf t for cuprous and cupric oxides (12.36 and 6.39, respectively) were used in the present work.

Jan 1, 1939

G. D. Kneip, Jr., and J. 0. Betterton, Jr. (Union Carbide & Carbon Corp., Oak Ridge, Tenn.)—The authors have contributed to the theory of zone melting by considering the effects of the solidification of the final zone on the distribution curves for the finite length bar. However, they do not consider in sufficient detail the restrictions imposed by the particular metals involved. Eq. 4 has the analytical solution which is equivalent to Pfann's' equation for normal solidification of the final zone. For distribution coefficients less than one, it is apparent from this form of the equation that, as the solidification approaches completion, the concentration attains extremely large values which are inconsistent with the density of physical materials. The necessary restrictions then are that the solute concentration cannot exceed the density of the alloy, or more frequently, the solubility limit in the solid phase. Secondly, the distribution coefficient cannot be constant from 0 to 100 pet solute unless the liquidus and solidus coincide over this whole region. It would thus be more correct to impose limits on the solubility such as would be indicated by a typical eutectic diagram. Furthermore, in this case the assumption that the distribution coefficient remains constant is more likely to be realized. Eq. 4 should then be replaced by the following C,,(x) =k/L-x [ ?11 3 C11 (x)Ddx - ?11 4 Cn(x)dx= k[Cn (L-l)/k] [ L-x/l]x-1 L-l = x = L-l [kCenterlie/Cn (L-l)]1/k-1 The typical shape for the distribution curve in the final zone is indicated schematically in Fig. 12 for the simple eutectic case. The solidification proceeds according to Pfann's' expression until the concentration reaches the maximum solubility in the solid phasc. At this point, the concentration changes abruptly to the eutectic composition and remains at this value for the duration of the solidification. The width of the flat region, and the back reflection of this effect into the concentration curve on subsequent passes, depends upon the ratio of the eutectic concentration to the original concentration, upon the distribution coefficient, and upon the number of passes. Similar effects on the concentration curves in the first zone length would be expected in many systems for which the distribution coefficient is greater than one, and where the maximum solubility is not too much larger than the original concentration. The discussers agree with the authors that the assumption that the liquid is uniform in concentration should be considered with caution. Kneip and Better-ton' have shown, however, that in floating zone refining of zirconium, using induction heating, the distribution of iron agrees with theory which assumes a uniform liquid composition. Hence, in this case, the experimentally realized distribution coefficient agrees with the present phase diagram within the experimental uncertainty. L. Burris, Jr., C. H. Stockman, and I. G. Dillon (authors' reply)—The authors are happy to have received the comments of G. D. Kneip, Jr., and J. 0. Betterton, Jr., which provide additional insight into the zone refining process. It is true that the equations developed are inapplicable if the solute solubility in the solid phase is exceeded. However, this restriction on the use of the equations was clearly stated in the paper. A closer approach to

Jan 1, 1957

By C. M. Wayman, R. Bullough

The selection rules fir twinning in fcc, bcc, and hcp lattices as established in Part I are used to predict the positions, relative to the matrix (untwinned) reciprocal lattice, of reflections due to double diffraction from the twins. These doubly diffracted reciprocal lattice points, which often coincide with ordinary matrix reciprocal lattice points, are deduced from kinematical considerations. 1. INTRODUCTION. A Bragg diffracted beam under certain conditions may be rediffracted to produce a third beam i .e., the incident beam, the first diffracted beam, and the rediffracted beam). In such a case an initial incident beam is diffracted by a set of planes; the diffracted beam may then be rediffracted by a second set of crystallographically different planes. The initially diffracted beam acts as an incident beam for the second set of planes, thus resulting in a twice-diffracted beam. This twice-diffracted beam may be regarded as a singly diffracted beam from a set of lattice planes whose structure factor is zero. In order to find the reciprocal lattice point corresponding to the doubly diffracted beam, it is customary1,2 to combine vectorially the component reciprocal lattice vectors which head to the extra reflections. For example, hcp metals may exhibit the forbidden (0001) reflections.* As Fig. 1 shows, the (0001) reciprocal lattice point results from: (1010) + (1011) = (0001) Similarly, combination of the beams (1011) and (1010) gives rise to: (1011)- (1010) = (0001) As Pashley and stowell2 have pointed out, even if in fact the double diffraction points lie on the Ewald sphere, they may not actually represent observed doubly diffracted beams because the initially diffracted primary beam, i.e., the source, must also correspond to a reciprocal lattice point which lies on the sphere of reflection; however, the reciprocal lattice point corresponding to the planes causing the rediffraction need not lie on the sphere. A little consideration of the Ewald construction in Fig. 2 makes this argument clear. In Fig. 2 the incident beam is represented by the wave vector ko. This beam is initially diffracted by the lattice planes (hkl) which are represented by the reciprocal lattice vector r(hk1) The reciprocal lattice point (hkl) of course lies on the Ewald sphere in order that the Bragg condition be satisfied. The initially diffracted beam is represented by the wave vector kl. If now kl is rediffracted by a second set of planes (h'k'l') resulting in k2 the lattice point (h"k"l") must lie on the Ewald sphere if the double diffracted beam k2 is to be observed. It is clear that (h"k"l") is the vector sum of (hkl) and (h'k'l'), and that (h'k'l') need not in fact lie on the Ewald sphere. In the usual1,2 analysis the origin (000) of the reciprocal lattice is shifted to the point (hkl). However the translation (000)— (hkl) automatically shifts the point (h'k'l') to the Ewald sphere, i.e., (h'k'l')— (h"k"1"). It is to be seen that (hkl), the source reflection, must lie on the Ewald sphere; if it does not the translation (000) — (hkl) places the "new" origin off the sphere, not satisfying the Ewald construction and hence the Bragg condition. The point (h"k"1") of course must lie on the sphere if it corresponds to an observed doubly diffracted beam. It is to be noted that crystals having the hcp structure, as just discussed, are intrinsically capable (i.e., without lattice defects such as twins) of double diffraction, as contrasted to fcc and bcc crystals which are extrinsically capable (requiring lattice defects or reoriented lattice regions). The "allowed" bcc reciprocal lattice points (which comprise an fcc lattice) are (110), (2001, (211), (220), (310), (2221, (321), (4001, and so forth. The combination

Jan 1, 1967

By Freeman Bishop

Having obviously cleared the way for fast confirmation by the Senate Interior Committee, the Administration recently named Elburt F. Osborn, vice president of Penn State University, as director of the Bureau of Mines. Interior Committee Chairman Henry M. Jackson (D of Wash.) has moved hearings on the nomination right along, forecasting that they will end in a confirmation. Dr. Osborn not only made a good impression on the Committee, he also made it amply clear that he would be the boss of his job and have right-hand aides to help straighten out conflicts stemming from recently enacted health and safety regulations for miners. Mining engineers in the field hope this clarification can come soon. They are frankly fed up with local, state, regional and national overlapping of restrictions.

Jan 1, 1970

By Kefton H. Teague

Olivine is a mineral containing a mixture of forsterite (Mg2SiO4) and fayalite (Fe2SiO4) in solid solution. The name olivine was first applied by Werner in 1790 (Hunter, 1941) because of the olive-green color of the mineral. Olivine is the principal component of the rock dunite, which itself is a member of the peridotite group of ultrabasic igneous rocks. European uses of the terms olivine and dunite are somewhat different than US usage. In commerce European usage defines dunite as containing 36 to 42% MgO, 36 to 39% SiO2, and loss on ignition of approximately 10%. The term olivine is used to designate a material containing approximately 85% forsterite, with a chemical composition of approximately 45 to 50% MgO, 40 to 43% Si02, 5 to 8% Fe2O3, and ignition loss of 1 to 2%. Uses The principal use of olivine is in various applications involving hot metal. Table I shows the best estimate available as to the amount of olivine consumed by various industries worldwide. In excess of 90% of world uses involve hot metal, with approximately 75% as a slag conditioner in the blast furnace production of pig iron. Approximately 15% of total olivine usage is as a special foundry sand in mold making for the brass, aluminum, magnesium, and manganese steel foundries (Schaller, 1957, 1958; Snyder, 1957; Anon., 1977). Olivine was first used as an industrial mineral in the early 1930s for a refractory material (Anon., 1970; Hunter, 1941). As a refractory raw material, olivine was first introduced in the United States as hand-cobbed, selected, shaped blocks of crude olivine. This use met with limited success. More recently, finely ground olivine blended with MgO and pressed into bricks, which are then fired, has found use in glass tank furnaces and open-hearth furnaces. Ramming or gunning mixes for basic furnace linings also utilize olivine. Olivine has been used in ladle linings with varied success. In Europe, substantial tonnages of olivine are utilized in refractory brick for night storage heaters (Anon., 1970). A limited amount of olivine has been used in the past as a fertilizer (magnesium source) and has been fused with rock phosphate to produce a magnesium phosphate as a plant food. The relatively high magnesia (MgO) content of olivine also attracted attention to it as a potential source of both magnesium compounds and as a source of metallic magnesium (Bengston, 1956; Hunter, 1941). The use of olivine as an additive in the blast furnace is relatively new, having been developed during the last decade. J.W. Currier (Private communication, 1978) describes the behavior of olivine in the blast furnace as follows: Olivine (Mg, Fe2, Si04) represents a unique lime-free source of magnesia and silica as a blast furnace charge component. Only recently has the magnesia-rich mineral been recognized as beneficial cleaner for blast furnace experiencing the problems associated with excessive alkali inputs and/or low coke stabilities. For some time it has been known that alkalis tend to recycle and accumulate in blast furnaces. In the high temperature region of volatilization the elemental alkalis formed attack the integrity of the graphite structure. This alkali attachment is enhanced when poor coke is used. Higher in the furnace stack volatile alkali cyanides condense and oxidize. This accumulation of alkalis tends to agglomerate the burden materials, producing scabs and scaffolds. These structures not only inhibit the downward movement of burden materials and the upward passage of wind, but also lead to erratic furnace performance

Jan 1, 1983

By G. R. Speich

The interlarnmelm spacing, growth rate, and degree of segregation that accompany cellular precipitation in four Fe-Zn alloys containing 9.7, 15.2, 23.5, and 30.5 at. pct Zn have been determined in the temperature range 400" to 600°C. The chemical free-energy change for the reaction was calculated from the available thermodynamic data and the known compositions of the phases. The fraction of the chemical free-energy change for equilibrium segregation that is converted into interfacial free energy decreases from 0.43 to 0.08 as the magnitude of this free-energy change increases from 35 to 270 cal per mole. At constant temperature the cellular growth rate is proportional to the cube of the dissipated free energy. At 600°C newly 100 pct of the equilibrium segregation is achieved during cellulm precipitation whereas at 400°C only 85 pct of the equilibrium segregation is attained. During cellular growth, mass transport of zinc occurs by grain boundary diffusion; excess zinc remaining in the a! phase after the completion of growth is removed slowly by volume diffusion. A modified Cahn theory of cellular precipitation predicts the observed interlamellar spacing within a factor of two. In cellular precipitation reactions such as pearlite formation or discontinuous precipitation, the basic problem is to predict the variation of growth rate G, interlamellar spacing S, and degree of segregation P with composition and temperature. To accomplish this we need three independent equations relating these quantities. One of these equations comes from the diffusion solution. To obtain two additional independent equations, some assumptions must be made. cahnl has suggested recently that two plausible assumptions are 1) that growth rate is proportional to the dissipated free energy and 2) that the spacing which occurs is that which maximizes the dissipated free energy. According to the first assumption, this spacing also maximizes the growth rate and the rate of decrease of free energy per unit area of cell boundary. The present work was undertaken to test these assumptions. To test the first assumption it is necessary to study a cellular reaction over a wide range of supersatura-tions to establish a relationship between G and the dissipated free energy at constant temperature. This is possible only in discontinuous precipitation reactions since in pearlite reactions constituents other than pearlite form if the composition of the parent phase deviates even slightly from the eutectoid composition. The Fe-Zn system was chosen for study because 1) discontinuous precipitation proceeds to completion over a wide temperature and concentration range, 2) the degree of segregation within the cell can be measured by lattice parameter measurements,2 and 3) the thermodynamics of this system have recently been determined by Wriedt.3 In this system the cells consisting d alternate lamellae of a and r phases form from supercooled iron-rich a phase. The a phase within the cells is bcc as is the original a phase, cia, but has a different orientation and a slightly lower zinc content than the original a phase. The r phase has a zinc content of about 70 at. pct and a crystal structure isomor-phous with T brass. EXPERIMENTAL PROCEDURE Four Fe-Zn alloys with 9.7, 15.2, 23.5, and 30.5 at. pct Zn were prepared from carbonyl-iron powder (400 mesh, 99.8 wt pct Fe) and zinc powder (200 mesh, 99.99 wt pct Zn). The powders were ball-milled together and cold-pressed under 60,000 psi to discs $ in. thick by $ in. diam. The cold-pressed discs of the alloys with 9.7 and 15.2 at. pct Zn were sealed in evacuated silica capsules and heated slowly to 1100°C over a period of 1 week (3 days at 600°C, then 3 days at 80O°C, then 1 day at 1100°C). The alloys with 23.5 and 30.5 at. pct Zn were treated similarly except that the final homogenization temperatures were 1000" and 85O°C, respectively, to prevent melting. The alloys were quenched in iced brine from the final homogenization temperature. Specimens of each alloy were subsequently aged in salt pots at temperatures of 400°, 450°, 500°, 550°, 600°, and 650°C for times that varied from a few minutes to several hundred hours. At a late stage of this work, an alloy containing 11.2 at. pct Zn was prepared by vapor-impregnation of iron foil with zinc vapor at 890°C. This alloy proved useful for electron microscope studies because it was free of porosity. The homogenization and aging conditions were based on the recent Fe-Zn phase diagram of Stadelmaier and Bridgers4 rather than the earlier diagram of ansen.5 They consist of a homogenization heat treatment in the homogeneous a field followed by an aging treatment in the two-phase a + r field. The aged specimens were metallographically polished and etched in 2 pct nital and the radius of the largest cell in the microstructure determined. This radius plotted vs time gave a straight line whose slope is the boundary migration rate or growth rate G of the cell. To determine the interlamellar spacing, specimens were examined by surface-replica and thin-section electron microscope techniques. Because of the irregular nature of the lamellae within the cell, the average interlamellar spacing S .of the cell was measured by the method of Cahn and Hagel,6 where S is defined by:

Jan 1, 1969

By B. F. Tillson, C. M. Haight

I. General Remarks..........................723 1. Location............................723 2. Characteristics of the Orebody..................725 (a) Mineralogical (b) Shape, Strike, Dip, Size, Etc. (1) Ore (2) Country Rock 3. Historical Sketch........................725 11. Mining Method..........................726 1. Outline of Methods Now in Operation............... 726 (a) Stopes (b) Fill . (c) Pillars (d) Slices (e) Transportation (f) Miscellaneous 111. Mining Operations........................727 1. Shaft..............................727 (a) Sinking (b) Dimensions and Equipment, Sketches (1) Size, Inclination. Direction, Etc. (2) Timbering (3) Stations (4) Track (5) Hoisting Equipment (a) Skips (b) Engines and Ropes (c) Idlers (d) Man Cars (6) Loading Pockets (a) Operating Figures (7) Headframe

Jan 1, 1918

By E. C. Burke, W. R. Hibbard

G. B. Craig (University of Toronto, Toronto, O,nt., Canada)—The advent of pyramidal slip at room temperatures in a magnesium crystal (M2,) and the very low value reported for the critical shear stress are puzzling. Table I1 shows the calculated values (approximate) for sin x cos I in this specimen if all the (10l) planes are considered. If So is the shear stress resolved in the direction of glide, on the glide plane, and U, is the stress operative in yield point tests, then: S,, — 325 (0.169) 55gpersqmm (approximately) for 1101: S,, = 325 (0.465) = 151 gpersqmm (approximately) for (0001): S,, — 325 (0.101) - 32.8 gpersqmm (approximately) Thus, we see the pyramidal plane in which the minimum shearing component operated. Barrett"' stated: "Experiments have repeatedly shown that, when there are several crystallographically equivalent slip systems in a crystal, the one having the greatest resolved shear stress will become active. ..." It does not appear to me that the measurements given are conclusive proof that pyramidal slip operated. E. C. Burke and W. R. Hibbard, Jr. (authors' reply) —Although the low value for the critical resolved shear stress for pyramidal slip may be puzzling, there does not appear to be any basis for questioning the validity of the measurements. Under the influence of constraints, slip systems having decidedly lower shearing components may become operative, e.g., "cross-slip" or the "forced glide" in magnesium reported by Bakarian. Since we believe that pyramidal slip occurred in the specimen because basal slip was inhibited, it is not too surprising to find an operative plane with a lower shear stress. -7 C. S. Barrett: Structure of Metals U94jl p. 2J6. New York. McGraw-Hill Book Co.

Jan 1, 1953

By L. Stone, H. Margolin

The Ti-C-N and Ti-C-O systems were investigated in the temperature range from 500° to 1400°C and in the composition range up to 2 pct C and 5 pct N or 0. Characteristic isothermal sections at 800°, 900°, 1000°, and 1300°C are presented. The Ti-N-0 system was studied in the temperature range from 900' to 1400°C with alloys containing up to 6 pct total alloying content. Characteristic isothermal sections at 1000° and 140O°C are presented. Melting-point data for all three systems are also included. THIS paper reports on one of a series of investi-gations which have been conducted on the phase diagrams resulting from interstitial alloying with iodide titanium. The other investigations involved delineation of the binary systems with carbon,' nitrogen and boron,h and oxygen. The Ti-0 binary system has also been investigated by Bumps et al.' In varying degrees, each of these interstitial elements has been shown to stabilize the low temperature a modification of titanium1-5 and each forms a face-centered cubic TiX compound (henceforth designated 6). In addition, the Ti-N and Ti-0 systems reveal a low temperature tetragonal phase (6) formed by a peritectoid reaction between a and TiX Experimental Procedure The development of experimental techniques for the study of titanium alloy systems has, to a large extent, become standardized. In this investigation, the equipment and procedures described in detail by Cadoff and Nielsenl have been used. Arc Melting: In general, binary alloys with carbon, nitrogen, and oxygen, prepared in the composition range of interest in this investigation, show negligible composition changes during arc melting. However, the possibility of the formation of some gaseous combination of alloying elements such as CO, CN, or NO during the preparation of these ternary alloys was considered. Calculations showed that the evolution of only 0.05 gram of such a gas would be detectable as a pressure change in the closed system used during preparation of these alloys. Such pressure changes were not observed. Consequently, nominal compositions have been used in plotting the data. The compositions of the materials used in the preparation of the alloys are shown in Table I. After melting for 3 to 5 min at 275 to 350 amp, the alloys were checked for homogeneity by microstruc-tural examination. Alloys containing up to 1 pct C were homogeneous in the presence of less than 3 pct N or 0. At higher alloying contents, some inhomo-geneities in the carbon distribution became evident. Alteration of the melting procedure toward longer times and higher currents did not improve the homogeneity of these alloys. Ti-N-0 alloys were homogeneous in the range to about 3 or 4 pct total alloying addition. Beyond this, almost all of the specimens showed as-cast microstructures consisting only of the phase. Consequently, inhomogene-ities could not be detected by examination of micro-structures. Ten alloys from each of the systems were analyzed for two of the elements present (oxygen being omitted in all cases and titanium being omitted in the Ti-C-N alloys). In all cases the analyses were found not to be sufficiently precise to serve as criteria for the total composition of these alloys. On the basis of phase distribution in heat-treated alloys, however, it appears that carbon is distributed throughout the alloys most uniformly, with oxygen and nitrogen following in that order. Heat Treatment: Specimens for heat treatment were wrapped in titanium sheet before sealing in the argon-filled quartz capsules. Heat-treatment times varied from 100 hr at 800°C to 0.5 hr at 1400 °C. After heat treatment the specimens were quenched by breaking the capsule in water. With the exception of alloys in the low composition region, heat treatment did not have an appreciable effect on the as-cast microstructures. Metallography: Following heat treatment, the specimens were prepared for metallographic examination by grinding on emery paper and electrolytic polishing. For the majority of the specimens a 10-sec etch with Remington "A" agent (25 pct HNO3, 25 pct Hf, and 50 pct glycerin) adequately

Jan 1, 1954

By H. L. Gilbert, W. W. Stephens

Production of anhydrous zirconium tetrachloride by direct chlor- ination of a zirconium oxide carbon mixture in a silica-brick-lined chlorinator is described. Theory and thermodynamics of reactions are discussed. A pilot-model chlorinator and full-scale production equipment are described and operating data are included. ANHYDROUS zirconium tetrachloride required as a starting material in the Kroll process for production of ductile zirconium has been produced in this country principally by chlorination of the carbide or "carbonitride" made by reduction of zircon sand concentrates with carbon in the arc furnace. This process and the equipment used have been fully described.'.' It is well known that zircon or badde-leyite ores may be chlorinated directly with chlorine in the presence of carbon, and this one-step approach would seem at first glance to be preferable to the two steps involved in the carbide-chloride operation. As previously discussed1 considerations leading to adoption of the longer process were: 1—Chlorination of the carbide proceeds rapidly at temperatures below 500°C, whereas temperatures above 900°C were considered necessary for chlorination of zircon-carbon mixtures. 2—The highly exothermic nature of the carbide-chlorine reaction makes it self-sustaining, while heat must be supplied continuously in direct chlorination of the ore. 3—Silicon was thought to be chlorinated along with zirconium in the direct chlorination of zircon, leading to high chlorine consumption. In production of carbide in the arc furnace, silicon is driven off as silicon monoxide and does not enter the chlorinator. 4—For efficient direct chlorination, the ore must be finely ground and intimately mixed with carbon, and the mixture briquetted. 5—Direct chlorination requires a much larger chlorinator for a given production capacity than chlorination of carbide. Recently it has become necessary to produce large quantities of zirconium metal from a chemically purified zirconium oxide. Since the cost of the latter is high, the relatively high losses encountered in production of carbide in the arc furnace could not be tolerated, and a graphite resistor furnace5 was developed for production of carbide, which was then chlorinated in equipment previously used for chlorinating arc-furnace carbide. This method of operation was quite satisfactory, and losses in the carbid-ing step were minimized. However, operating costs were relatively high, and the process did not lend itself particularly well to large-scale operation because of the multiplicity of small units required and the hand labor needed to load, unload, and maintain the furnaces. To chlorinate the carbide, a vertical-shaft chlorinator was used in which the charge was heated with a central split graphite-rod resistor.' Operation of this chlorinator was somewhat less satisfactory with the resistor furnace carbide than it had been with the arc-furnace carbide due, principally, to the differences in physical properties of the carbide. The arc-furnace carbide is obtained as a fused metallic-appearing mass which can be crushed to -1/4 in. with production of a minimum of fines, whereas the resistor-furnace product is lightly sintered and produces a large proportion of fines in crushing and handling. These fines tend to pack in the chlorinator and promote channeling. which results in poor chlorine efficiency and low capacity. Data based on production of 22,000 lb of chloride from resistor-furnace carbide in this equipment are shown in Table I. Necessity for increasing production of chloride from about 2,000 to 15,000 lb per week led to further investigation of possible methods for direct chlorination of the oxide or oxide-carbon mixtures. Direct chlorination of the pure oxide presents much less difficulty than chlorination of zircon sand or oxide ores. The oxide is easily ground to —200 mesh, and no silica is present to cause excessive consumption of chlorine or contamination of the product. Theoretical Considerations The principal chemical reactions involved in chlorination of mixtures of zirconium oxide and carbon may be written as follows: ½ ZrO2 (c) + C(c) + Cl2(g) = ½ ZrCl,(g) + CO(g) [I] 1/2 ZrO, (c) + CO(g) + Cl2(g) = ½ ZrCl,(g) + CO2(g) 121 ½ ZrO2 (c) + ½ C(C) + Cl2(g) = 1/2 ZrCl4(g) + ½ CO2(g) [3]

Jan 1, 1953