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By J. B. Hess, R. L. Dietrich

In the mechanical twinning of magnesium on the {1012} planes the crystal-lographic deformation is such that, in the direction of the hexagonal axis [0001], twinning is possible only undef tension stress, while in the directions perpendicular to the hexagonal axis, twinning is possible only under applied compression. Thus in extruded magnesium and magnesium alloys where the extruded textures are best described as either [1010] or [1120] fibers,' the basal planes are all nearly parallel to the extrusion axis and such material is susceptible to ( 1or2 ) twinning by compression in the direction of the extrusion axis. In order to explain the results of cyclic stressing in alternate tension and compression upon the tangent elastic modulus in such material, Dorn and Thomsen2 have postulated that, following twinning under applied compression, a retwinning or un-twinning occurs as a result of residual microtensile stresses immediately upon removal of the applied compressive force. Recently, Carapella and Shaw4 have used this supposition of untwinning in magnesium in discussing their results on indentation hardness and on cold drawing of MI sheet. In trying to confirm this hypothesis by suitable direct experiments, the authors have been unsuccessful in detecting any evidence whatsoever that untwinning follows the release of applied compression and are forced to conclude that this hypothesis is improbable or, at least, that the quantity of material involved is small. For the experiments, small cylindrical specimens 0.040-in. diam. X 0.120- to 0.140-in. long were etched from 0.0625-in. extruded rod. The ends of the specimens were accurately faced off parallel in a jeweler's lathe. The compression jig was simply an ordinary set of micrometer calipers with an anvil interposed between the spindle and the specimen. The anvil was free to move axially but restrained from rotating so that the rotation of the spindle would not introduce a twisting stress on the specimen. With this arrangement the compressive strains could be estimated to 0.0001 in. and X ray pinhole photographs could be taken with or without load as required. Fig I is the pattern of an uncompressed specimen of AZ3I alloy, showing the approximate [1010] extrusion fiber. The X ray beam is normal to the extrusion axis; in this position it is particularly sensitive in detecting twinning if one directs his attention to the diffraction ring from the basal plane. Since twinning on (1072 planes produced an 86" reorientation of the crystal lattice, the original basal reflections occur at the 3 and 9 o'clock positions while the (0002) reflections from any twinned material will occur at the 6 and 12 o'clock positions. Fig 2 shows the same specimen compressed 3.0 pct, photographed while still under load. This strain has been just enough to produce a detectible amount of twinning and the sensitivity is roughly indicated by the fact that no alterations were produced in the intensities or maxima of the other diffraction rings. A strain of

Jan 1, 1949

By R. K. Matheson

THE Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 19o tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65° From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. CHOICE OF RAISE A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. TIME SCHEDULE As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1942

By H. S. Cannon, F. N. Rhines

The application of a static load to a copper powder compact during sintering at an elevated temperature accelerates the rate of sintering in such a way that a given load induces the same proportional increase in rate for all times of sintering. It is shown that sintering under a load is like creep under a fixed load in that the stress required to accomplish a given degree of densification is proportional to the logarithm of the sintering time. VIRTUALLY all of the theories of sintering that have been put forward within recent years have contained the assumption that the chief driving force of the process is the surface tension resident in the exposed surfaces and internal pores of the compact, or powder mass. An externally applied load might be expected to provide a somewhat equivalent driving force for sintering. It is known that the application of a compressive load, during sintering, hastens densification, but it is by no means clear whether sintering under the influence of surface tension alone and sintering under an applied load are fundamentally similar processes. The present study was undertaken in an effort to obtain an answer to this question and with the hope that the answer may contribute to an understanding of the mechanism of sintering. It has been found that there is a close relationship between sintering and creep processes. Copper powder compacts were sintered in hydrogen at 1000 °C under dead loads of 0 to 165 psi and the progress of sintering was observed by means of density measurements. Using a commercial reduced oxide copper powder, see Table I, cylindrical compacts Yz in. x x in. high were made at a pressure of 12,500 psi. For sintering, these were placed in a cylindrical graphite container into which they fitted snugly. A compressive axial load was applied to the compact through a graphite rod of similar diameter, which rested upon the upper end face of the compact; iron weights, in suitable amount, were placed upon the graphite rod to provide the load. This assembly was encased in a vertical silica tube within a resistance-type furnace capable of maintaining a substantially constant temperature within the working zone. After various predetermined times at temperature, the samples were removed from the furnace and their density was measured by conventional means. The heating-up and cooling-down times have been included in the computed sintering time, using the assumption that the sintering rate doubles for each increase of 10 °C in temperature. The experimental results are presented in Table 11. Upon metallographic examination of the specimens, it was found that there was no apparent change in the shape of the pores as a result of loading, i.e, no flattening, Fig. 1. Substantial contraction away from the container walls was observed in all samples sintered with loads of 10.5 psi and less.

Jan 1, 1952

By R. F. Ashton, D. S. Thompson

In reporting results of precipitation hardening experiments, it is customary to include such conditions as solution heat-treatment temperature, specimen size, and quench medium as well as the aging time and temperature. One portion of thermal history that is often overlooked is the effect of the heating rate employed to achieve the aging temperature. It is the intent of this note to demonstrate the importance of the rate of heating to aging temperatures. This work was carried out on the aluminum alloy 7075 aged at 162oC, a temperature commonly used to obtain stress corrosion resistant tempers. A 2 in. thick plate of 7075 was quenched into 24°C water after solution heat-treatment for 2 hr at 470°C. Approximately 13 in. around the periphery of the plate was discarded to minimize quench rate variations, and the remainder was divided into 0.5 by 1.8 by 2.0 in. blocks which were held at room temperature, 24oC, for 4 days. This aging, or incubation period, at room temperature allowed a copious supply of GP-zones to form. Blocks were then heated at various rates to 162oC using the practices shown in Table I. Both single- and two-step aging practices were employed for each heating rate. The single step consisted of heating directly to temperature at the indicated rate while the two-stage practice included an interruption of heating for 7 hr at 107°C. In the case of the two-stage 100°C per min heating rate experiment, the specimens were returned to room temperature after 7 hr at 107°C where they were held for 16 hr be-fore finally heating at this rate to 162°C. This devia-tion in practice is not considered to be significant. Blocks were removed after various times at 162oC, and aging curves were determined from duplicate R. F. ASHTON and D. S. THOMPSON,Member AIME, are Senior Scientists, Metallurgical Research Division, Reynolds Metals Company, Richmond. Va. Manuscript submitted December 2. 1968. IMD Table I. Aging Practices Two-step Single-Step Heating Rate* Aging Aging Heating Method oC per Min A B Programmed Air Oven 0.23 * 0.01 C D Programmed Air Oven 0.93 ? 0.03 E F Air Oven at Temp -10 G H Oil Bath at Temp ˜100 *Heating rate was measured with a thermocouple at the center of a specimen. An average value between the starting temperature and 3°C below the aging temperature is reported for practices E-H. PRACTICES 70 — A B C E and G 50 - 2 4 6 810 20 40 60 HRS.AT 162°C Fig. 1—The influence of processing variables on 7075 aging curves. Fig. 2—7075 aged according to practice H (single-step aged, fast heating rate) for 3 hr. Magnification 100,000 times.

Jan 1, 1970

By Fred W. Stiefel

THIS paper covers the operation and maintenance of Diesel locomotives underground on a portion of the Delaware River Aqueduct.[t] This part of the tunnel is 15 miles long, with shafts 14 ft. in diameter spaced about 5 miles apart. These shafts (2, 2a and 3) are 825 ft., 1551 and 840 ft. deep, respectively. The grade of the tunnel is almost level, except for the 12,100 ft. north of shaft 2, which has a grade of 2.31 per cent. The tunnel is circular, 17 to 19 ft. in diameter. It lies at a depth of 400 ft. below sea level, with a maximum over-burden of 2400 feet. The tunnel driving was carried on in six headings, using 15 storage-battery locomotives, each of 13 gross tons and having a maximum speed of 8 miles per hour. The longest haul was 3 miles, which proved to be the maximum distance for which storage-battery locomotives could be employed economically for the amount of material to be moved. One locomotive was used for switching cars at each heading, one for switching cars at the bottom of each shaft and one on the main-line haul from the heading to the shaft. The maximum amount of material hoisted in one shaft from two headings when 104 ft. of tunnel was driven in 24 hr. was 1000 cu. yd., or about 2000 tons of rock in place. The total quantity excavated was 750,000 cu. yd. (1,500,000 tons of rock). The tunnel-lining operations using concrete presented a problem of hauling a maximum distance of 5 miles in the tunnel. The concrete-placing equipment had a capacity of about 2000 cu. yd., or 4000 tons in 24 hours. Experience had shown that storage-battery equipment could not handle this quantity of material within the time required at the distance of 5 miles. Trolley-locomotive haulage was considered dangerous and expensive, and would cause interference because transmission equipment would be required. Therefore the performance of Diesel locomotives of 160 hp., with a speed of 18 miles per hour, in comparison with that of other types of locomotives, was studied. CHOICE OF LOCOMOTIVES In deciding upon the type of locomotive to be used for the 5-mile haul, four alternatives were considered, as follows: I. To purchase additional storage-battery locomotives with the necessary batteries over the 6 locomotives and 12 batteries available. 2. To convert the existing battery locomotives to the trolley type. 3. To purchase new trolley locomotives. 4. To purchase Diesel locomotives. PROPOSITION No. I The 90-hp. storage-battery locomotives available (each weighing about 12 tons), when hauling a load of 64 tons, with a drawbar pull of about 1900 lb. on a level track, developed a speed of 8 miles per hour and could make five round trips in 8 hr. with a

Jan 1, 1942

By H. Landsberg

THE curriculum presented here is an outgrowth of discussions by the Committee on Geophysics Courses of the A.I.M.E. in previous years. It had to be a compromise between the desires voiced by employers of geophysicists, professional geophysicists and the possibilities at an average college. There one has to face the questions of the available teaching staff, the inertia of administrators and teachers set in their ways, and the red tape involved in modernizing antiquated curricula. It is, of course, not ideal to have colleges and universities lag 10 years behind the times. Rather, instead of being prodded by industry, -they should be ahead of industrial needs. There should also be a certain amount of leeway for the student in any college curriculum. All these points had to be considered in drawing up a curriculum, which, instead of being narrowly termed "geology" or "geophysics" was rather called "earth science." It gives the student a chance to specialize in one of the various branches represented by teachers on the faculty of The Pennsylvania State College, giving as options Geology (with subdivisions of Mineralogy and Paleontology), Geophysics, Geography, and Meteorology. Table IA. represents the option of geophysics in the curriculum that envisages that a student may want to make his career in any of the fields encompassed by the American Geophysical Union or in geoexploration. This is, of course, only the start of an education that in most cases will have to be followed by graduate work for complete professional training. The outlined curriculum is definitely of an engineering type and is identical for all of the four options mentioned for the first two years. Specialization will start in the junior year but even thereafter the branches are kept as closely allied as feasible in order to permit a student to change from one to the other (e.g., because of a different aspect in the employment situation) with a minimum of additional work. Table IB shows the geophysics option according to subject matter, indicating the considerable amount of work in the fundamentals of mathematics, physics and chemistry with a liberal number of geology courses, and some engineering. The courses in the field of geophysics are limited-just two lecture courses as a general introduction and one field course, to permit the student a visualization of the correlation between mathematically analyzed physical data and the geological realities. Thesis work is considered essential to teach the student, while he is still in college, some things S. F. Kelly has mentioned as desirable: (I) following of instructions for somewhat independent work; (2) organizing of material, and (3) producing a readable report. This, together with the personal contact with the advising faculty member, will help in forming a student's personality. The seminar course attempts to get the student into the habit of reading current literature, so that he can keep abreast of the times.

Jan 1, 1941

By R. A. Stauffer

THE chief ore of lithium is spodumene, a lithium-aluminum silicate containing up to 3 pct lithium. The preparation of lithium salts from spodumene is costly because of the low concentration of the metal and the difficulty of separating it from the other alkali metals present. Dr. Kroll's and Dr. Schlecten's recent paper' has reviewed various methods proposed for preparation of lithium. It will be seen from this paper that the methods proposed to date have required one or another of the lithium salts as a raw material. The high cost of these salts has been the chief item of expense in the lithium metal produced. This situation has made it interesting to investigate the possibility of producing the metal directly from spodumene. When spodumene is calcined at 1000°C in the presence of lime, lithium oxide is replaced in the ore as is evidenced by the fact that it may be leached from the reaction mixture with water.2 Free energy calculations suggested that lithium oxide might be reduced with aluminum or silicon and the lithium produced distilled from the reaction mixture by heating to about 1000°C in a vacuum. These facts taken together suggested that it might be possible to prepare lithium metal directly from spodumene by heating a mixture of the ore, lime, and a suitable reducing agent in vacuum and distilling off the lithium metal as formed. THE REPLACEMENT OF Li20 IN SPODUMENE A series of experiments were undertaken to determine the ratio of lime to spodumene necessary to free substantially all the lithium oxide present in the ore. The spodumene used in these experiments was supplied by Foote Mineral Co. in the form of powder of about 200 mesh. Analysis of the spodumene was as follows: [ ] Lime (cp -200 mesh) was added in various ratios to spodumene and thoroughly mixed in a ball mill. Two hundred gram samples of spodumene-lime mixtures were charged loosely in fireclay crucibles and held at 1000°C in a gas-fired furnace for 5 hrs. The product in each case was ground to pass 140 mesh and then placed in a thimble in a soxhlet extractor. Extraction with water was carried on for 6 hrs. After this period the extract was reduced in volume by evaporation in order to precipitate dissolved lime and was then filtered and titrated with 1/4 normal hydrochloric acid. After correcting for residual calcium oxide in solution, the alkali present was calculated as lithium oxide. The results of these experiments are shown in Table I and Fig I. At lime-spodumene ratios less than 1.5, the products were invariably fused. At the ratio of 1.5 and above, the product was loosely sintered with little evidence of fusion. When the ratio of lime to spodumene was three, replacement of alkali oxides in the

Jan 1, 1947

By J. W. Borough

THE electrical conductivity of pure copper is markedly decreased by small amounts of impurity in solution. The magnitude of this effect has been very carefully determined by numerous investigators who studied a wide variety of impurity elements. Much of this work is summarized in an excellent review by Pawlek and Reichel.' The case of zinc impurity in copper is an exception to the above in that there has been no well-documented study of the effect using high-purity starting materials. The values for the effect of zinc on conductivity, which are usually taken from unpublished work by Skowronki, indicate that zinc in low concentration has no effect whatever in lowering the conductivity of pure copper. This effect, if true, is unusual since it is generally known that even in lowest concentration, impurities exert an effect on conductivity which is proportional to the amount dissolved. It is more reasonable to suspect that in the original work on the effect of zinc some of the zinc was oxidized out of solution during specimen preparation. The following experi- and calculated d-spacings based on the unit cell dimensions of Baenziger et al. Since the single crystal studies of Baenziger et al. were limited to (hk0) and (hk 1) data, it was not possible to Compare all the intensities of the powder and single crystal patterns. However, the strongest reflections of both studies, the (321) reflection, agree very well when the appropriate multiplicity factor is used. Good agreement was also found in a number of other reflections of lesser intensity. Despite the precautions taken in annealing the U6Fe powder, an appreciable amount of UO, was detected in the powder pattern after heat treatment. The powder pattern reported by Groganoand Clews is given in Table I. Except for the 3.15A line, which is the strongest line of UO,, their pattern, as far as it goes, compares favoEably with the data from this study. Gordon and Kaufmann have reported a diffraction powder pattern purported to be U, Fe but it bears very little resemblance to the patterns of Table I. ments were carried out to redetermine the effect of zinc on conductivity using high-purity starting materials and processing conditions capable of producing oxygen-free alloys. Materials used in the experiment were ASR 99.999 pct Cu and Cumingo 99.999 pct Zn. Four 1 lb melts were prepared with identical procedure in argon atmosphere with crucibles made from Speer No. 580 graphite. The compositions were 0.0, 0.1, 0.3 and 1.0 pct Zn. Standard No. 14 wire (0.064 in. diam) was made from these melts as follows. The 1 pct Zn sample was first reduced 94 pct by cold swaging with five annealing treatments of 800o for 10 min in hydrogen between swagings, and then reduced 94 pct to the

Jan 1, 1962

By G. S. Scott, G. W. Jones

ECONOMIC waste caused by mine fires may become considerable,15 especially if a fire is allowed to spread or temporarily to get beyond control. It is important, therefore, to act promptly whenever a fire is suspected and to make every action count. To contribute toward its extinction effectively it is necessary to know at the earliest possible moment, and continuously thereafter, what is happening at the seat of the fire, which may be in a remote or inaccessible area, and also to recognize what may be expected as a result of any proposed combative measure. In combatting a fire the mining engineer and the chemist should cooperate, first deciding upon the course of action to be taken and then drawing upon the mining engineer's intimate knowledge of the physical structure of the fire area, the cost and time required to carry out proposed operations, and the probable efficiency with which these operations could be performed, and upon the chemist's knowledge for the results to be expected therefrom. After the initial steps to be taken have been decided, it devolves upon the mining engineer to see that they are carried out and upon the chemist to evaluate the results and to keep the mining engineer continuously informed in regard to a variety of matters-such as the composition and toxic or asphyxiating properties of the atmosphere in which the men may be working, whether the fire is increasing or decreasing in intensity and its probable temperature, the quantity of air leakage and the explosion hazards of the fire area. The purpose of this paper is to discuss what the chemist can and cannot contribute as a member of the fire-fighting personnel. The technical literature contains several hundred papers relating directly and indirectly to mine fires, and this fact alone is testimony to the complex nature of the phenomena involved in the birth, life, and final demise of a fire. A fire may be stopped (I) by preventing oxygen from reaching the coal or (2) by removing the conditions necessary for the rapid reaction of coal with oxygen; i.e., decreasing the temperature. All methods of prevention or control. necessarily must be based upon one or the other, or both, of these two principles. . The chemistry of a fire may be understood readily if the low-temperature oxidation of coal is considered as a starting point; then oxidation at higher temperatures (or combustion) ; and with these, the composition of the solid and gaseous products resulting from the oxidation or from the effects of the heat liberated. This discussion refers particularly to Pennsylvania anthracite, although much of the material applies equally well to coals of lower rank. OXIDATION OF COAL Oxidation of coal fundamentally is confined to the surface; that is, the rate at which oxygen penetrates into the interior of a solid block of coal is so slow that it may be ignored.6,16 Oxidation starts when a fresh surface is exposed, and the rate is

Jan 1, 1942

By G. A. Koehler

The Hiawatha mine, a producer of iron ore, is near the City of Iron River, in Iron County, Michigan. This mine, operated by The M. A. Hanna Co., has been chosen by the author as a typical iron mine in this district that is using the cut-and-fill mining system. General Geology The Hiawatha ore body lies along the steeply dipping north limb of a syncline that closes around on the east end to form a bathtub-shaped structure. Throughout this major structure there are very many smaller wrinkles or folds, which vary the direction as well as the thickness of the ore body. The total length of the ore formation with minor interruptions so far developed is about 3600 ft. The ore body throughout the entire present depth of 2000 ft. has averaged approximately 40 ft. in thickness; it is somewhat narrower at the upper levels and shows a marked increase in thickness on the bottom levels. This increase in thickness probably indicates that the bottom of the trough is near or that a major monoclinal fold has been encountered, which flattens the ore and gives apparent horizontal thickening. The footwall of this ore body is generally a black, graphitic slate, very rich in iron sulphides, quite combustible when exposed to air, and very treacherous to handle during mining operations. The hanging wall, in most places, is a cherty, fairly hard iron formation, and in general stands quite well ill open stopes. There are places, however, where black slate is found in both the hanging wall and the footwall, which makes mining difficult, as the black slate tends to slough off after being exposed to air. In these places it becomes necessary to leave a scale of ore 5 to 6 ft. thick on both hanging wall and footwall, which is not recovered. Fig. I shows a typical vertical cross section through the Hiawatha ore body. Type of Stoping Because of the size and shape of the ore body and the hard, dense structure of the ore, sublevel stoping has been an efficient and economical method of extracting the ore. This method, with various minor changes, has been used since the mine was opened in 1893. There have never been any stope-filling operations, however, until recent years, when crushing of pillars and black-slate fires forced the issue and methods were arrived at to fill all stopes. Development of Deeper Levels Shaft sinking to deeper levels in development operations has been done successfully by two methods without interference or hindrance to production. These systems are: I. Sinking an auxiliary winze a short distance from the shaft, then drifting back to the shaft location and putting up a pilot raise through to the bottom of the shaft. This raise is always in the manway compartment of the shaft, so that no work is ever done under a cage or skip. A 20-ft.

Jan 1, 1946

By P. W. Barton, A. A. Johnson, E. J. Hughes

THE only systematic work on the effect of carbon on the recovery and recrystallization kinetics of iron so far reported in the literature is that of Venturello et al.1 These workers found that doping high-purity iron so that it contained up to 86 ppm C caused a marked retardation of recovery but had little or no effect on the recrystallization kinetics. Such evidence for an effect as was obtained was inconclusive. In the course of a study of electron-beam zone-refined iron the authors have recently examined the effect of added carbon on the recrystallization kinetics of this material. The amount of carbon added was about 50 pct more than was used by Venturello et 01. and definite evidence was obtained that carbon at this level retards recrystallization. The starting material was taken from a batch of 0.2-in.-diameter Johnson-Matthey spectroscopically pure iron rods which had been further purified by being given three passes at 4 in. hr-1 in a standard MRC electron-beam zone-refining unit. The purity of this material. which was used in two other investigations of the properties of high-purity iron,2,3 was established by an extensive series of analyses. Four laboratories, the National Research Corp., Newton, Mass., the Naval Research Laboratory, Washington, D.C.. the Battelle Memorial Institute, Columbus, Ohio, and the National Physical Laboratory. Teddington. Middlesex, England. carried out analyses for carbon. oxygen. and nitrogen. and in addition a complete mass-spectrograph analysis was carried out at Battelle. In all, twelve analyses were carried out for carbon, seventeen for oxygen and twelve for nitrogen. The amounts of these elements found to be present were 10 * 7, 8 i 5, and 0.9 i 0.1 ppm, respectively. The limits given are rms deviations. Two results, one of 46 ppm of oxygen and another of 31 ppm of nitrogen, were rejected as anomalous. The mass-spectrograph analysis revealed the major impurity to be 25 ppm of tungsten. All of the other elements detected were present in amounts less than 0.5 ppm. Some of the rods were doped with carbon by being immersed in a suspension of carbon in oil, heated in a vacuum at 400°C, and then zone-leveled to make the distribution of carbon uniform. Carbon analyses carried out on three of the doped rods gave a mean value of 130 ± 45 ppm. Some of the doped and undoped rods were deformed 58 pct by swaging and cut into specimens 1/2 in. in length. Isothermal recrystallization experiments were then carried out on them at 575", 600°, 615", 630°, and 650°C, the volume fraction of each specimen recrystallized being measured by a point-counting method described by Hilliard and Cahn.4 The results for the two types of specimens are shown separately in Figs. 1 and 2. In Fig. 3 the pairs of recrystallization curves for each temper-

Jan 1, 1965

By S. C. Sircar, B. K. Dhindaw

KINETICS and mechanism of sulfation reactions have been studied extensively. Qualitative and quantitative data are reported for the kinetics and mechanism of sulfation reaction of copper oxide pellets;''2 yet it remains a controversial Subject. The mechanism and kinetics of the sulfation of zinc oxide powder (B.D.H. analar quality) were studied by gravimetric technique. A spring of high-carbon steel wire with accuracy well within 0.2 mg was used for measuring weight change. The sample was taken in a Pyrex glass bucket, hanging freely inside a Pyrex glass tube reactor, with a gold wire attached to the spring, the reactor being surrounded by a movable furnace (double wound type) with a temperature control within *5"C. Flow of oxygen, 60 cu cm per min, and sulfur dioxide, 45 cu cm per min, were measured through calibrated flowmeters and the mixture dried over PZ05 before entering the reactor. The experimental results are presented in Figs. 1 and 2. The plot of weight change of zinc oxide, giving the amount of zinc sulfate formed, was found to be linear over the temperature range of 500" to 700° C and up to a conversion of approximately 60 pct, as beyond it they divert from linearity.' The rate of sulfation was found to be affected by total flow rate of gases at a fixed ratio of SO2/O2, up to 60 cu cm per min, and was found to be linearly dependent on quantity of total powdered sample taken. The plot of rate of sulfation of zinc oxide, 70 mg, with varying amounts of vanadium pentoxide, at a constant ratio of gases as mentioned above, and at a constant temperature of 625"C, was also found to be linear as seen in Fig. 2. Extrapolation to zero V2O5 addition gives the same rate of sulfation, as in case of pure zinc oxide, 70 mg. Vz05 was prepared by heating ammonium vanadate at 600°C for 15 min in a muffle furnace. The available literature points out that the theory of sulfation that has received the maximum support postulates the following mechanism for sulfation of ZnS: 2ZnS + 30, = 2Zn0 + 2SOZ; the ZnO formed acts as a catalyst for the reaction, 2 S02 + O2 = 2S03. This SO3 reacts with ZnO to give zinc sulfate. It is suggested here that the above widely supported mechanism is operative in the sulfation reaction and that the rate-controlling step in the sulfation reaction of ZnO is the formation of SO3 from SO, and 0,. The authors are carrying on further work in this line, and the details will be published later. The authors wish to acknowledge the facilities provided and the help extended by Prof. P. R. Dhar, Head, Department of Metallurgical Engineering, and the permission granted to publish this note by the Director, I. I. T., Kharagpur.

Jan 1, 1969

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

Discussion of the paper of G. P. Hulst, presented at the Salt Lake meeting, August, 1914, and printed in Bulletin No. 92, August, 1914, pp. 1865 to 1871. L. S. AUSTIN, Salt Lake City, Utah.-One question that I would like to ask is in regard to the sampling kettle which stands beside the softening furnace. How is the sample taken or melted? G. P. HULST.-The sampling kettles in question are of 45 tons capacity and are fired with coal. The lead bullion comes in on a high line in car-load lots of 30 to 40 tons. The entire lot is trucked from car, weighed over scales and thrown into the sampling kettle, which has been previously cleaned. The kettle is fired and the bullion melted, then heated up until no frozen bullion' remains in the .bottom or adheres to the sides of the kettle. The kettle is scraped clean, sides and bottom, and the wet dross is skimmed Doff clean into molds. The kettle is stirred thoroughly during the period in which 24 to 30 samples of approximately ½ assay ton are taken. These samples are taken in a long-hand loci iron mold, providing for taking six to eight "gumdrops." The mold is inserted and heated to the same temperature as the molten metal, and gumdrops are clipped out and cooled in the mold by dipping 'the bottom of. the mold into Water. The samples should be free from fins. The gumdrops represent the contents of the kettle, and in assaying, they should not be clipped, but weighed up and the result computed. The dross bars are weighed and sawed according to old hammer and punch sample template. The sawdust represents a sample of the dross. In eases of high gold in the bullion there is a slight correction in the assay. This method of sampling was developed at the Omaha plant of the American Smelting & Refining Co., W. T.. Page, Manager, in 1903 or 1904, for adjusting differences in checking smelter sampling. L. S. AUSTIN.-The base bullion which you treat is very clean. It has been remelted, and skimmed before shipment. How rapidly can you get it through the softening furnace, and how does it compare with ordinary base bullion? G. P. HULST.-Twenty-four hours is the softening limit on these charges. In other words, it is charged into the furnace, heated up, and the first and second skimmings are taken off, and it goes into the kettle the next day; that is, 24 hr. elapse from the time the bullion is charged to the time it is in the kettle. L. S. AUSTIN.-It seems to be a long time to hold such a pure bullion. I suppose, though, it comes around conveniently so that you have a time for all the other operations of refining. Isn't that the case? MR. HULST.-Every refinery is based on those principles.

Jan 11, 1914

By William Loblowitz

A simplified system of flotation testing has been developed recently at the Bureau of Mineral Research of Rutgers University. The technique involves the known principle of skin flotation and can be used with minute quantities of sample and reagents in a beaker. A vessel containing the conditioned sample with added water and flotation reagents is held inclined and rotated slowly. Water-repellent particles come in contact with air and float on the surface of the liquid, while wettable particles are submerged. The float is poured off and filtered whereas the nonfloat remains in the beaker. The filtrate may be returned to repeat the process. This procedure can be made quantitative. Separations are obtained and are remarkably complete and repeatable. The effect of different reagents, optimum quantities, pH and other variables can be determined in a fraction of the time required for conventional testing. However, the method gives no information on the effect of frothers and the quality of froth to be expected in a cell, and it is not practical for particles finer than 200-mesh. Unlike other preliminary tests, separate float and nonfloat products are obtained, and their contents can be determined quantitatively by chemical analysis or grain counts. After the optimum reagent combination has been determined, the effect of fines and frothing may be verified in the conventional flotation cell. On the other hand, agglomeration-tabling can be directly duplicated from beaker tests. Procedure: Five to ten grams of sample are weighed into a 100-ml beaker. Flotation reagents and a small initial quantity of water are stirred together; additional water is then added to produce the proper water-solids ratio. The beaker is then tilted to 45" and rotated slowly. After three to five rotations, liquid and float are poured off and filtered; the filtrate may be recycled until the separation is completed. One test generally requires less than 10 min. Reagent concentration and pH remain quite constant because of the recycling of the filtrate. Two to four tests may be made simultaneously, and the effect of varied conditions may be observed by comparison. Considerable time is saved in preparation and handling since small quantities of sample suffice for a series of systematic tests. Repeatability One test was repeated four times under constant conditions. The sample used was a quartz sand containing approximately 24 pct feldspar and some clay aggregates. The +28-mesh grains had been screened out, and most of those finer than 100-mesh had been removed by hydraulic classification. The results are presented in Table I. For comparison, 500 g of the sample were separated by agglomeration-tabling using the same reagent combination and quantities as given in Table I. The weight percent float was 18.3 pct, composed almost entirely of feldspar and with only traces of quartz. Most of the remaining feldspar was recovered in a middling product with 2.2 pct by weight.

Jan 1, 1952

By E. H. Burdick

The difficulties inherent in table concentration operations as applied to gold, silver, lead and zinc ores, are accentuated in the scheelite mill, which has a flowsheet that is similar in general principles. The narrow "spread" in the specific gravity between scheelite and the usual gangue in which it occurs often makes its clean separation extremely difficult and in some cases impracticable. The problem is further complicated by the extreme friability of the mineral and its inordinate tendency to "slime," particularly where fine grinding is necessary and wet grinding methods are employed. The physical aspect of the occurrence of scheelite in the ores, the kind and character of the rock in which the mineral is present, its hardness, friability and the specific gravity of its constituents, as well as like characteristics of the associated minerals— pyrite, pyrrhotite, sphalerite, and others, all have a bearing on the successful milling of crude ores and the production of a concentrate that is marketable with minimum price penalty or none at all. Most tungsten deposits are considered low grade when the limiting factor of low percentage recovery by gravity methods is borne in mind. The great majority of deposits developed or in process of development carry from 0.4 to 1.0 per cent tungstic oxide. Ore carrying I per cent is considered good grade; 0.7 to 0.8 per cent, fair, and below that figure, low grade. Steady mill feed averaging upward of I .o per cent WO3 is exceptional. Inasmuch as buying specifications demand a 60 per cent W03 content, the mineral content of the ores as mined must be concentrated to at least that degree. A few companies have constructed chemical plants for the treatment of their complex concentrates and low-grade flotation product, but the smaller operator must rely on table concentration, borrowing the tools and to a large extent the methods of his predecessors in the gold, silver, lead and zinc mills. Whether these are the best tools and methods that can be devised is an open question. Except in rare instances, table concentration, however carefully conducted, does not yield a satisfactory final recovery of the mineral content in headings as indicated by chemical determinations of tungstic oxide content. A saving of 60 to 65 per cent is considered fair, and an average 80 per cent recovery, exceptional. Each scheelite ore body is a separate problem. Details of handling that give best results in one case may not be successfully operative at a property on the same contact zone and but a short distance removed. Supplementary treatment of tailings by flotation may be successful in producing a low-grade concentrate of around 15 per cent W03, which is adapted to further treatment chemically at a given property and entirely unsuccessful at another. This problem constitutes another of the unanswered questions. The following paragraphs relate solely to the beneficiation of scheelite ores by table concentration and supplementary methods.

Jan 1, 1943

By E. H. Burdick

The difficulties inherent in table concentration operations as applied to gold, silver, lead and zinc ores, are accentuated in the scheelite mill, which has a flowsheet that is similar in general principles. The narrow "spread" in the specific gravity between scheelite and the usual gangue in which it occurs often makes its clean separation extremely difficult and in some cases impracticable. The problem is further complicated by the extreme friability of the mineral and its inordinate tendency to "slime," particularly where fine grinding is necessary and wet grinding methods are employed. The physical aspect of the occurrence of scheelite in the ores, the kind and character of the rock in which the mineral is present, its hardness, friability and the specific gravity of its constituents, as well as like characteristics of the associated minerals— pyrite, pyrrhotite, sphalerite, and others, all have a bearing on the successful milling of crude ores and the production of a concentrate that is marketable with minimum price penalty or none at all. Most tungsten deposits are considered low grade when the limiting factor of low percentage recovery by gravity methods is borne in mind. The great majority of deposits developed or in process of development carry from 0.4 to 1.0 per cent tungstic oxide. Ore carrying I per cent is considered good grade; 0.7 to 0.8 per cent, fair, and below that figure, low grade. Steady mill feed averaging upward of I .o per cent WO3 is exceptional. Inasmuch as buying specifications demand a 60 per cent W03 content, the mineral content of the ores as mined must be concentrated to at least that degree. A few companies have constructed chemical plants for the treatment of their complex concentrates and low-grade flotation product, but the smaller operator must rely on table concentration, borrowing the tools and to a large extent the methods of his predecessors in the gold, silver, lead and zinc mills. Whether these are the best tools and methods that can be devised is an open question. Except in rare instances, table concentration, however carefully conducted, does not yield a satisfactory final recovery of the mineral content in headings as indicated by chemical determinations of tungstic oxide content. A saving of 60 to 65 per cent is considered fair, and an average 80 per cent recovery, exceptional. Each scheelite ore body is a separate problem. Details of handling that give best results in one case may not be successfully operative at a property on the same contact zone and but a short distance removed. Supplementary treatment of tailings by flotation may be successful in producing a low-grade concentrate of around 15 per cent W03, which is adapted to further treatment chemically at a given property and entirely unsuccessful at another. This problem constitutes another of the unanswered questions. The following paragraphs relate solely to the beneficiation of scheelite ores by table concentration and supplementary methods.

Jan 1, 1943

By R. A. Vandermeer, P. Gordon

The study of recrystallization in cold-worked metals is complicated by the impossibility of making direct three-dimensional optical measurements. Metallo-graphically, theinternal three-dimensional features of samples must be inferred from their intersections with the two-dimensional plane of polish. The inherent difficulties involved in this procedure have led— particularly during the last few years—to some instances of rather doubtful interpretations of experimental data in terms of nucleation and growth rates in recrystallization. In the present paper an analysis ; is given of the recrystallization characteristics in a polycrystalline alloy containing 0.008 wt pct copper in ZONE-refined aluminum *and an effort is madf to reemphasize the necessity in such studies for attention to the proper statistical relationships between one- two- and three-dimensional features of opaque samples. It is shown that, in common with several recently described cases for other types of transformations,1-5 recrystallization in this alloy can best be rationalized on the basis that all nucleation took place essentially at time zero—i.e., there is no significant time-dependent nucleation during annealing—and that the kinetics are controlled exclusively by the growth-rate parameters. This is true despite the fact that the number of

Jan 1, 1960

THE Institute of Metals Division of the A. I. M. E. has joined with the American Society for Steel Treating and the American Welding Society in support of National Metal Week in Philadelphia, Oct. 8 to 12. Individual and joint technical sessions, plant inspection trips, and a general exposition to be held in Commercial Museum have all been arranged. With the coming of the Institute of Metals Division into the circle, the scope of the exposition has been broadened to cover non-ferrous metals, and exhibits of much interest are promised. Headquarters for the Institute of Metals and the American Society for Steel Treating will be at the Benjamin Franklin Hotel, and for the American Welding Society at the Bellevue-Stratford. The American Society for Steel Treating and the Welding Society will hold technical sessions each forenoon at their hotel headquarters. The Institute of Metals Division will hold sessions Tuesday morning and Wednesday and Thursday afternoons; that for Wednesday being a joint session with the American Society for Steel Treating. The latter society will hold independent afternoon sessions on Thursday and Friday. Registration and information desks will be open at the hotels throughout the week, and an attractive program of entertainments has been arranged.

Jan 1, 1928