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By Frank G. Breyer

C ONDITIONS have not been favorable for new developments in any line. It has been a period, how- ever in which recent developments have been subjected to the severest tests. Those which have been able to hold their own in the severe competition of 1932 must be considered fully proved commercial processes. The continuous vertical-muffle retort plant at Palmer- ton is reported to have been in operation at practically full capacity throughout the year. The high quality of metal produced and the low operating costs are factors which have kept this plant running when many of the older furnaces were shut down. The plant at Meadow- brook, W. Va., is also reported to have been in steady operation throughout the year. Average retort life is said to be better than two years with about three-quarters of the original retorts still in operation after 2 ½ years.

Jan 1, 1933

By M. M. Rao, P. G. Winchell

The growth rates of bainitic plates were measured at 400°C in Fe-Ni-C alloys containing 0.10 atom-fract~on nickel and 0.0012 to 0.0075 atonz-fraction carbon. The growth rates are adequately represented by where xc is nearly the atom fraction carbon of the bulk austenite and PXCy is nearly the carbon atom fractlon in the ferrlte of radiids p In equilibrium with austetzite. The form of the equation is that predicted by a model in which carbon diffusion in austernite controls the gvowth, but the numerical constatnt is two orders of magnitude below that suggested by the model. THE growth of bainitic plates in steel is often assumed to be controlled by the diffusion of carbon away from the advancing plate tip. This hypothesis predicts that the growth rate will increase as the carbon content of the austenite, xCz, is reduced toward the carbon content of the saturated ferrite comprising the plate tip, PxCY The growth rate should vary approximately as (xCg- pxCy)-1. Experimental observation of the growth behavior at low carbon levels should provide a significant test of this model. An alloying element in addition to carbon is required so that low-carbon austenite can be experimentally observed while undergoing bainitic transformation. Nickel was selected. The presence of nickel complicates the interpretation of the data in two ways: First, diffusion of nickel during the transformation would make analysis very difficult. Nickel is assumed immobile during the transformation. Second, nickel affects the solubility of carbon in ferrite and austenite in equilibrium. This effect has been evaluated.' At the completion of our experimental work Goode-now et al.2 published data in essential agreement with the observations to be reported here. Since their discussion is abbreviated and their data are scanty in the region of interest, we believe the present work is of significance. I) THE MODEL OF BAINITIC PLATE GROWTH The rate of lengthening of a plate is assumed to be controlled by the diffusion of carbon from the advancing ferrite-austenite interface into the surrounding austenite. The precipitation of carbides is assumed to be a secondary process. For ease of analysis the carbon-atom ratio,* pxCy, of austenite in equilibrium with ferrite which is convex with minimum curvature radius p, and the carbon-atom ratio, PxCY, of that ferrite in equilibrium with austenite are assumed independent of location on the ferrite-austenite interface. Since these carbon contents vary with the radius of curvature of the ferrite, p, their assumed positional independence must be held as an approximation. The consequences of these assumptions have been developed approximately by zener3 and Hillert,4 and the resulting equation for a platelet has been applied to bainite by Speich and cohen5 and Kaufman, Radcliffe, and Cohen.8 The Zener-Hillert equation* for plates is: The analysis of Hillert is supported by that of Hor-vay and cahn7 which involves no mathematical approximations but does include the assumption that the a/y interface coincides with an isoconcentration line. The solutions of Horvay and Cahn for elliptic paraboloids are replotted in Fig. 1. The shape of the paraboloid is expressed in terms of the ratio of the principal radii of curvature at its tip, A =p1/p2, which is also the ratio of the minor to the major axis of the elliptic cross section. The Zener-Hillert equation for plates is also plotted. The agreement is within a factor of two for (pxyaCr - xyC )/(xyC - PxCaY) between 0.5 and 100. This is the range of interest here and in most other work on bainite. The original form of the Zener-Hillert equation was the form given above with the right-hand side replaced by (pxCya -xCy)/(PxCya). This replacement is not appropriate here. 11) THE EXPERIMENTAL PROCEDURE Alloys were prepared and three kinds of experiments carried out. Continuous-cooling-transformation experiments were carried out on wires by measuring temperature and resistance during continuous cooling. Isothermal-transformation experiments were carried out on wires by measuring electrical resistance as a

Jan 1, 1968

By A. W. Schlechten, Ernest J. Breton Jr.

Experiments on the separation of copper and zinc ions by selective action of ion exchange resins showed the carboxylic type to be more effective than the sulphonic resins. The latter demonstrated a greater capacity over a wider pH range. Data show the effectiveness of resins as a means of concentration. IN recent years the restrictions of stream pollution laws and the high price of metals have created an interest in ion exchange as a means for metal recovery. Some applications have proved successful. In Germany during World War 11, 17 tons of copper per day were recovered from rayon mill wastes by means of ion exchange resins;' and for some time in this country a large ion exchange unit has been in operation for the recovery of copper from rayon waste water. The possibilities of applying ion exchange to the recovery of metals occurring in plating rinse water is particularly promising. In most of these applications only the metal being recovered occurs in the waste. The ion exchange resins act merely as a means of concentrating the metals to a point where they can be recirculated. It would be highly desirable to use ion exchange as a means of not only concentrating but also of separating metals. With the exception of the impressive separations accomplished in connection with the atomic energy program, very little has been done on metal separations.' Therefore, an investigation was undertaken at the Missouri School of Mines and Metallurgy to determine if either of the two main types of ion exchange resins could be used to separate metal ions in solution. The selective removal of copper ions from a mixture of copper and zinc on carboxylic and sulphonic-type resins was investigated as a function of flow rate, pH, copper-zinc ratio, and concentration. It was shown that zinc can be separated from copper and that very large ratios of concentration can be obtained using ion exchange resins. Since ion exchange is relatively new to the field of metallurgy, a brief review of the subject will be included. Theory of Ion Exchange A comprehensive theory for ion exchange has not been developed as yet, but the mechanisms are analogous to metathetical reactions: R Na + Cu++ *=? K(SO3)2 Cu + 2Na+ R is the designation for the ion exchange resin. If a copper solution is passed over a resin bed in the sodium form, two ions of sodium will be released for every ion of copper removed. For the most part this reaction follows the laws of mass action and of electrical neutrality. Consequently, if an excess of sodium ions is passed over a bed containing copper, the reactions will be reversed, and the resin will be regenerated to its original form. A few empirical rules governing the exchange reaction have been set forth: 1—In general ions with a high valence will replace ions with a lower valence. 2—Ions having higher activity coefficients have a higher replacement potential. 3—In a series of mono-valent ions, those with the smallest radii of hydra-tion will tend to replace those having larger radii of hydration. 4—Where ions are similar in most respects, those with the higher atomic weight sometimes will take precedence. This last rule is not as definite as some of the others. These rules apply to rather dilute solutions at moderate temperatures and assume all ions to be present in about equal concentrations. Higher concentrations and temperatures may in some cases reverse the normal exchange reactions. Ion exchange materials are unique in that their efficiency increases as the concentration of the solution decreases. For many exchangers, most efficient operation is obtained at concentrations in the order of one thousandths of a percent. Most applications, though, are made in solutions containing considerably higher concentrations than this. Coste9 as shown that ion exchange resins will remove aluminum and iron effectively' from solutions of up to 10 pct chromic acid. Ion Exchange Resins Ion exchange resins are insoluble, porous, resinous structures to which active groups have been attached. Active groups such as (—SO,,)- and (COO)- pick up cations; hence structures saturated with groups such as these are called cation exchangers. Structures saturated with groups such as (—NH,)' which pick up anions, are referred to as anion exchangers. The resinous structure of necessity is resistant to strong acids, bases, oxidizing, and reducing agents, and most of the common organic solvents. An idea of the stability can be gaged from the fact that resins last for many years under constant use without detectable chemical or physical breakdown. The ion exchange reaction is not confined to the surface of these synthetic resins. Its porous structure permits active groups in the center of a particle as well as those on the surface to remove ions. A high capacity resin such as Amberlite IR-120 will remove up to 3.3 lb Cu per cu ft of resin. In this investigation several approaches to the problem of separating copper from zinc by ion exchange were considered. First, if a reagent could be found which would complex one of these metals and not the other, then by passing this reagent through a bed of exchanger containing copper and zinc, the

Jan 1, 1952

By R. J. Hopkins, L. B. Haney

On the basis of limited experimental work conducted at the Port Pirie smelter, it would appear that, by increasing the specific surface of sinter, and possibly that of coke as well, a marked increase in the rate and degree of sinter reduction in the furnace shaft can be achieved. Fusion-point tests have shown that this increased reduction means a much higher sinter fusion point and narrower plastic zone, while the more efficient utilization of coke in the furnace is inherent in the intensification of reduction by the provision of a more reactive sinter surface. Investigations aimed at increasing the ability to operate the blast furnace with a higher lead tenor of sinter reducing fuel cost proportionately, and at eliminating the loss in reduction potential in coke entering the furnace, as shown by the presence of appreciable proportions of CO in the top gases, are reported. Several lines for future investigations that could result in improved smelting methods are proposed. A SOMEWHAT informal presentation of ideas on lead blast-furnace smelting will be given in this paper. These ideas have been developed during observations of furnace behavior at Port Pirie and earlier at Mt. Isa, and have been guided by the results of laboratory experiments at Port Pirie. The authors do not intend to propound any particularly new theories, but present their ideas as a basis for discussion which may help toward a better understanding and improved control of blast-furnace operation. Performance variations of considerable magnitude occur from time to time at most lead blast-furnace plants. While Port Pirie is fortunate in having a very regular supply of high grade concentrates, it is no exception in this respect. Such irregularities are of considerable economic importance and they constitute a challenge to present-day metallurgists in their efforts to obtain an ever greater degree of control over furnace operation. In operating as Port Pirie does with a relatively high lead charge, it has been recognized for a long time that shaft conditions are ,of the greatest importance. A practical, or engineering, attack on the problem of accretions consisted of changes in furnace design, e.g., by widening the shaft to 10 ft above the bottom row of tuyeres, and by the provision of curved ends and vertical walls.' That this attack has had a great measure of success is shown by the fact that barring down of these wide furnaces is rarely necessary, shaft conditions have greatly improved, and sensitivity to charge variations has been markedly reduced. These modifications, followed by some metallurgical changes (chiefly the incorporation of pig iron into the furnace charge in 1948), have brought furnace operation to a point where some variables which were previously masked can now be studied. Accordingly, over different periods of time during the last five years, such variables as could be given numerical values have been recorded and subjected to statistical examination. The objectives of this work were: 1—to reveal the true importance of the different variables and 2—to show the proportion of total furnace variation which is unexplained and therefore referable to causes not measured at present? e.g., shaft accretions, sinter structure, coke structure, and others. Statistical Study of Variables In this work factors relating to furnace speed and lead reduction have been examined: 1—slag com-

Jan 1, 1955

By R. W. Smith, K. H. Lee

Bond type simulated closed-circuit grindability tests and batch grindability tests were run on six pure minerals plus a limestone and a taconite ore. The simulated closed-circuit tests were made at five different meshes of grind. The minerals tested varied widely in hardness and breakage pattern. For all simulated closed-circuit tests both Bond and Holmes work indices were calculated and compared. The Bond grindability testlm6 has been widely and quite successfully used for predictions of such items as ball mill power requirements and as an aid in plant comminution equipment selection. The test is interesting in that although run as a series of batch tests it is over-all a simulated closed-circuit test with a circulating load of 250%. The simulated closed-circuit feature is probably the primary reason for the notable success of the test. Perhaps the major criticism that can be brought against the test is that it is time consuming, since it normally takes a trained technician about an eight-hour shift to perform it.4 For many purposes this is an acceptable time requirement. However, if a grindability test is to be used for mill control purposes it should be performed in as short a time interval as possible, preferably under an hour. Sometimes in industry strictly batch tumbling mill tests are run and an attempt is made to correlate such batch tests with plant mill operation. Then an attempt is made to adjust operating tumbling mills based on data from the tests. The tests, however, frequently do not correlate well with operating data. The purpose of the present work was to explore some of the similarities and differences between a simulated closed-circui.t grindability test and a batch type test. Six different minerals plus a limestone and a taconite ore were tested in the experimentation. These materials varied widely in hardness, breakage characteristics, etc. Standard dry Bond grindability tests were mn on all materials at 48, 65, 100, 150 and 200 meshes of grind. Feed material was prepared by stage cmshing to -6 mesh. Initial feed volume was 700 cc. Care was taken to maintain the standard constant weight (about 20125 gm) and number of balls (283) during all the tests. Rotational speed was 70 rpm. Table I lists the materials studied along with pertinent data regarding the materials and Bond grindabilities obtained. In addition, dry batch tests were performed on all materials. These tests were made in the same mill as was used for the Bond tests. An identical ball charge and an identical feed volume were used. After rotating the mill for a predetermined number of revolutions at the same rotational speed as for the Bond test, the material was removed and separated from the ball charge. A screen analysis was performed on this product. Knowing the original size distribution of the feed and the size distribution of the product, it was possible to calculate the grindability or grams of a particular mesh produced per revolution of the mill. After the screen analysis of the product, the various fractions were recombined and placed back in the mill for further grinding. After grinding for additional revolutions of the mill the procedure was repeated. Thus, it was possible to obtain grindabilities for each material at a number of meshes of grind and for various revolutions of the mill. Grind-abilities obtained from batch tests are shown in Table 11. COMPARISON OF BOND AND BATCH GRINDABILITIES Figs. 1A through 1D show relationships between batch and Bond grindabilities. Figs. lA, lB, 1C and 1D show respectively graphs indicating the relationships for batch tests run for 100, 300, 500 and 700 revolutions of the mill. For a particular material on each graph, points are shown for 200, 150, 100, 65, and 48 meshes of grind except for several missing coarse meshes of grind for galena which did not fall conveniently on the figures. For each material and graph the point nearest the origin is for 200 mesh, the next for 150 mesh, the next for 100 mesh, etc. Considering Fig. lA, it is interesting to note that all relationships shown are essentially straight lines. There are, however, considerable differences

Jan 1, 1969

By H. E. Lee, P. C. Feddersen

Greenockite is the only known cadmium mineral of importance. It occurs rather universally, in minor concentrations, as a secondary mineral in sphalerite deposits. The world's cadmium output is obtained through the processing of metallurgical by-products, largely from the treatment of residues from electrolytic zinc, retort zinc and lithopone plants. These sources are supplemented by the processing of fumes from lead and copper smelting operations. The development of modern selective flotation practice in the decade 1920-1930, which permitted the economical mining of complex lead-zinc ores, resulted in significant increases in the quantities of cadmium entering lead smelting systems. Being closely related to zinc as to occurrence, properties and production, most detailed description of cadmium recovery methods are to be found recorded in connection with zinc metallurgy. Other than occasional articles pertaining to particular operational procedures, literature offers but little in the nature of a balanced survey of lead smelter cadmium recovery practices. While many of the basic operations described for the recovery of cadmium from zinc by-products are applicable to the treatment of lead plant products, the inherent problems involved differ widely. In general the cadmium content of related by-products from routine lead smelting operations is present in lower concentrations, exists in a less soluble state and is associated with both a greater quantity and a greater variety of detrimental impurities. To cope with these problems, lead smelter practices are found to follow the general outline: Preparatory Processing 1. Concentration operations 2. Sulphation operations Cadmium Plant Processing 1. Leaching operations 2. Purification operations 3. Sponge precipitation operations 4. Metal recovery operations 5. Refining and casting operations As in the case of related zinc plant operations, cadmium recovery practices at lead smelters are not standardized. They not only vary as to type, but also extent. Depending upon prevailing conditions, lead smelter cadmium operations range from simple concentration campaigns, for the purpose of sufficiently "up-grading" products for shipment elsewhere, to complete processing steps for the production of refined metal. Preparatory Processing The cadmium content of lead smelter receipts is low and, as a rule, proportionate to the zinc content; the usual range of cadmium contained being of the order of 0.01-0.05 pct. Were it not for the low boiling point of cadmium, such small concentrations would, no doubt, be lost in the large tonnages of slag, metal and other smelter end products. However, the ready volatility of cadmium and its compounds at prevailing lead smelting temperatures results in its concentration in fractional portions of fume collected. This collected fume comprises a circulating load within the smelter system. Thus, the cadmium content of blast furnace fume* increases with each successive circulation until an equilibrium value is reached when the sum of the cadmium losses, due to handling and in slag, waste gases and other end products, becomes equal to the intake as ore. With ore receipts averaging, say 0.03 pct cadmium, the concentration value obtainable, through fume circulation in a routine manner, approaches 10-12 pct. In such operations, cadmium concentrations in blast furnace fume of from 3-5 pct are readily .attainable. However, the concentration gain beyond this range, with each additional circulation, is progressively decreased as a result of mounting losses occurring through handling and in end products. Therefore, to avoid excessive cadmium loss and to enhance the ultimate concentration anttainable, it is customary practice to isolate blast furnace fume at some intermediate cadmium content for special concentration procedure. The most common type cadmium concentration "campaign" involves special smelting operations wherein a relatively high portion of blast furnace fume at 3—6 pct cadmium is incorporated into the sinter charge. This type practice is roughly illustrated by the diagram on p. 111. Cadmium fume collected from lead blast furnace operations is not amen-

Jan 1, 1950

By Eugene F. Broome, Hugh A. Walls

Self-diffusion coefficients were measured using the modified shear cell technique over a temperature range from 31" to 401°C. These data agree within experimental error with those of Petit and Nachtrieb, which were determined from 30°to 99°C. LIQUID self-diffusion behavior in gallium has been studied previously by Petit and Nachtrieb&apos; using their shear cell technique.&apos; Data were obtained in that investigation over temperatures ranging from 30.6 to 98.7"C, and these comprise the only published data for liquid gallium. In the work reported here, diffusion data were obtained from 30.8 to 400-9.Cusing a shear cell technique different from that of Petit and Nachtrieb. This modified shear cell method and the mathematical analysis for diffusion in a finite capillary geometry have been previously described.3 PROCEDURES AND ANALYSIS The segments of the modified shear cell were fabricated from types 304 and 316 stainless steel. No corrosion problems with gallium were encountered until attempts were made to obtain data at 500°C. The diffusion capillary diameter was 1.59 mm and capillary lengths for the four sets of cell segments used are given in Table I. Gallium with a purity of 99.999 wt pct was obtained from Var-Lac-Oid Chemical Co. The gallium-72 isotope used to measure the self-diffusion coefficients was produced from this inert gallium by irradiation in The University of Texas Reactor Facility, and was checked for radiochemical purity. Following a period of isothermal diffusion controlled to within 0.0l°C, the contents of each capillary section were emptied into a counting planchet, weighed to assure that gallium had not been lost during handling, and the radioactivity determined by windowless gas-flow proportional counting. Great care was taken to obtain a reproducible counting geometry for the samples. Diffusion coefficients were obtained from the solu-tion3 to the diffusion equation for the modified shear cell problem, viz.: in which B2= C2(t)Li + L2)- CSL1-COL2 P* (C0-Cs)(L1+ L2) and L, and L2 are the capillary segment lengths. The step-function initial concentration distribution was expressed as C(x,O) = C, for -L1 = x < 0 and C(x,O) = Co for 0 < x 5 L?. Also, C2(0 = ^- fcix,t)dx Computation of the diffusion coefficients was done numerically as described in our earlier paper.3 Because of the short half-life of gallium-72 (14.1 hr) and the relatively long counting times taken to give a precise value of the diffusion coefficient, it was necessary to account for the radioactive decay both during the actual count and between counts for different samples (before and after diffusion). Thus, the average count rate Cz(t) for a second sample counted from times t3 to t4 can be made consistent with the average rate Co determined for a first sample counted from times t1 to t2 by the relationship: N(e-»i-e-M») , , C2(f) " &-«,)(*-«.-«-»«) [21 in which N is the total number of counts taken and is the decay constant. RESULTS AND DISCUSSION The self-diffusion coefficients determined in this investigation are listed in Table 11. Values of the parameters used to calculate the coefficients are given in the table, as are values of the random counting error &apos; for the concentration ratio pz and the corresponding expected random variation AD in the diffusion coefficient. The quantity is calculable from: &apos;Cg — C2(t)} , „ ,2 .-„¦, M co-csl(ACs)J W in which AC is the error in the individual count rate. The mean values of the diffusion coefficients within each temperature group are also given in Table 11. Such temperature groupings of data are reasonable because of the small variations between the temperature~.~

Jan 1, 1970

By J. L. Smith, A. L. Albee

This article presents a series of guidelines by which the geologist can evaluate the likelihood of surface faulting and its probable extent at any given site in Southern California and Nevada. The information is intended primarily for geologists concerned with establishing design criteria for proposed nuclear plants. The geologic problems involved in the location of a nuclear power plant are fundamentally no different from those for other types of installations. They fall into four main categories: foundation stability, landslides and slope stability, shaking due to earthquakes, and surface faulting. For problems in the first three categories the foundation engineer, the geologist, and the seismologist can provide criteria for plant location and design, and these problems can generally be economically handled by appropriate design measures for a project of the magnitude of a nuclear plant. However, problems in the last category — surface faulting — are more difficult to handle and require a unique evaluation. Accordingly, this paper will deal primarily with the problem of establishing design criteria for surface faulting, particularly as it affects nuclear facilities. A nuclear reactor is a power source that for greater safety is contained in a heavy, air-tight structure, just as gas, oil, water, and other power sources must be contained. Surface faulting is significant in that it may reduce the integrity of the containment by affecting critical exterior piping or by breaching of the containment. A similar significance exists relative to dams or tanks for the storage of water, gas, or oil, except that in these latter examples the breaching of the container automatically releases the fluids to do their damage. This is not necessarily the case with the rupture of a reactor containment structure because the function of the containment is totally protective, i.e., it is necessary only in the event that radioactive products are released from the reactor, and there are many other safeguards to pre- vent this release even if the containment is ruptured. At the present time, nuclear power plants must be located near large sources of water for cooling the steam generated. The construction of an industrial facility of any kind on the coast line is esthetically distasteful to most people since, unfortunately, there is not enough coast to fulfill all the needs and all the desires of all the people. In most cases where industrial facilities encroach on the lives of citizens, there is no mechanism other than zoning laws by which a person can effectively protest. But in the case of a nuclear facility, the public hearing required by the Atomic Energy Commission provides a forum for dissent, as in recent case histories, and the question of safety provides objectors with a weapon for fighting the construction of the plant. The nature of a public hearing for a nuclear plant is such that the prospective owner and operator must prove that there is no undue hazard, whereas the objectors need only demonstrate that there is a reasonable doubt. It is in this situation that geology becomes the Achilles Heel of nuclear power plant location. For his investigation, the geologist has natural exposures of rock at the ground surface and a limited number of trenches and drill holes to give him a fairly complete picture of the distribution of the various rock types. From the surficial data, he must infer three more dimensions — depth below the surface, past time, and future time. Unlike many problems faced by engineers, the geologist has only this one set of data from which to reach a conclusion — he is unable to reproduce the natural sequence of events in order to obtain another set of data for comparison. Hence, by the very deductive nature of a geologic conclusion, it is difficult to prove a geologic conclusion beyond a reasonable doubt even to other geologists — and perhaps one should say especially to other geologists because the experience and background of a geologist will strongly influence his conclusion, and no two geologists have exactly the same experience and training. The engineer and the public official would like the geologist to conclude that faulting cannot occur at a given site or to assign a numerical value to the probability of its occurrence — but no responsible geologist can do either of these things. Since government officials and others must make decisions that affect the public safety, it would seem that the geologic profession must attempt to establish criteria and

Jan 1, 1968

By Frederick Forscher

Frederick Forscher (Nuclear Materials and Equipment Gorp.)—It is gratifying to see another series of experimental results that essentially reconfirms our reported observations of "Strain-Induced Hydrogen Embrittlement in Zirconium" (F. Forscher: AIME Trans., 1956, vol. 206, p. 536, noted as Ref. 3 of paper). However, several important considerations may have been overlooked, or—at least—have not been specifically pointed out. It is the purpose of this discussion to present some of these considerations. The statement is made that "the equilibrium concentration of hydrogen in solution at 20°C in zirconium is approximately 10 ppm; i.e., no zirconium hydride precipitates are formed at or below this hydrogen content." This is perhaps the case in the unstressed condition. In the presence of triaxial (plastic) strain however, even as little as 1.4 ppm H2 (Forscher&apos;s Figs. 6 and 11) will cause porosity near fracture in 20°C tests. The necked-down region that develops during a tensile test provides the triaxiality necessary for these pores to open. Both Beevers&apos; and Forscher&apos;s micrographs show that porosity is concentrated in the necked region. So-called "pure zirconium" containing about 10 ppm H2 also shows pronounced porosity near fracture (Beevers&apos; Fig. 3). Since it is generally agreed that such porosity is

Jan 1, 1965

By J. M. W. Mackenzie

The rheological properties of aqueous solutions of several commercial flocculants were investigated using a rotating coaxial cylinder viscometer. Similarities between the pH-apparent viscosity relationship and the flocculating properties of the reagents are discussed here. Synthetic neutral polymers and polyelectrolytes are used extensively as flocculants in mining and chemical engineering operations. The action of these reagents has been interpreted as adsorption of the polymer flocculant to the mineral surface, followed by the formation of physical polymer bridges between the particles.4,5 A consequence of the bridging model is that the efficiency of a polymer flocculant depends on the degree with which the polymer is adsorbed to the mineral surface, and the extent to which the polymer molecule forms a bridge between colliding particles. The length of the polymer molecule and its orientation at the mineral surface are important in the bridging mechanism since the further the polymer segments can protrude from the mineral surface the greater the probability may be of these segments contacting another mineral particle. Linke and Booth7 have shown that, for a silica suspension, both the settling rate at a given polymer addition and the maximum settling rate obtainable increase with increasing molecular weight. The analysis of silberbergl0 has

Jan 1, 1964

By Klaus-Peter M. Hanke

INTRODUCTION In early 1979 Jim Walter Resources, Inc. (JWR) of Brookwood, Alabama approached TMCI Construction, Inc (TMCI) to make a proposal on a program that involved the sinking of up to 10 ventilation shafts of approximately 6.7 m (22 ft) diameter and ranging in depth from 500 to 700 m (1650 to 2300 ft) for the JWR coal mines in Alabama. At this time TMCI was already constructing the first spiral underground bunker (capacity 2000 tons) in North America for the JWR organization at their No. 4 mine in Alabama. The TMCI proposal was based on the use of the mos modern large diameter shaft boring machine rather than sinking the shafts using the conventional drill blast-muck technique. The proposal was made based o: the experiences by the parent company, Thyssen Schachtbau, which has been using this type of machin in Germany for shaft boring since 1971. As a result of the TMCI proposal JWR issued a purchase order to TMCI for the construction of four 6.7 m (22 ft) diameter, concrete lined, unfurnished ventilation shafts ranging in depth from 500 to 700 (1650 to 2300 ft). An order was thus placed with WIRTH Machinen- and Bohrgeraete Fabrik GmbH, in Germany for the manufacture of a model 650/850 E/Sch "Schachtbohrmaschine" (Vertical Shaft Borer = V-mole which arrived on site in Alabama in early 1981. The first V-mole GSB 450/500 was introduced in Germany in 1971 and was capable of enlarging in one step a pre-drilled 1.2 m (4 ft) pilot hole to 4.5 - 5.0 m (14.7 - 16.4 ft). This machine has sunk 9 staple shafts and deepened one surface shaft for a total of 2360 m (7740 ft) of shafts. On the last shaft boring operation in 1978 the machine was converted as an experiment to drill without a pilot hol using a hydraulic pumping system to remove the cutting debris. A second generation machine, the SB VI 500/650, was introduced in 1977 for enlarging the pilot hole to a range of 5.0 - 6.5 m (16.4 - 21.3 ft) diameter. This machine is still in operation and has already drilled well over 2000 m (6500 ft) of shaft. The third generation of V-mole, the SB VII 650/85( for diameters from 6.5 to 8.5 m (21.3 to 27.9 ft) was: commissioned in May 1980 and has been used for two surface shaft deepenings totalling 606 m (1990 ft) with another scheduled for 1982. The main advantages favouring the use of such V-moles were identified as: 1) A reduction in manpower to the crew required in a conventional shaft sinking operation. 2) A considerable reduction in time to complete a shaft compared to conventional techniques. 3) The use of the V-mole eliminates many of the hazards encountered in conventional sinking. Based on the successful performance of the first three V-moles in Germany, Thyssen Schachtbau decided to employ this principle abroad. In 1980 a second machine of the third generation was built and is now operated by TMCI Construction, Inc. in Alabama. The first shaft was completed at the end of 1981 and this paper describes the method of operation including some unique aspects not attempted on prior V-mole operations and some of the statistics arising out of the experiences during the first shaft boring operation. THE NO. 7 MINE FAN SHAFT SITE Jim Walter Resources, Inc. was formed in 1970 to exploit the coal field in Alabama on the southern tip of the Appalachian coal field. The coal reserves amount to around 650 million tons of mainly good quality coking coal of which about 350 million tons are to be extracted over the next 30 years. Shaft sinking and preparatory work began in 1972, and at present 6 mines are producing around 5.4 million t.p.a. Annual production is to expand to 10 million t.p.a. as soon as possible, and the ventilation shafts to be sunk by TMCI play a vital role towards attaining this goal. The first shaft site is located at the No.7 mine, near Brookwood, Alabama. The actual location of the shaft relative to the production shafts is shown on the mine plan (Fig. 1), which also shows the room-and-pillar extraction system used at present. The mine plan further shows the conveyor route used for the muck removal. The geological survey showed that the strata consisted of horizontal layers of mainly sandstone, sandy shale and shale interspersed with several coal seams. The seam being extracted at the No. 7 mine is a combined seam made up of the Blue Creek and Mary Lee seams at a depth of 513 m (1682 ft) and having an average seam thickness of about 2 m (6 ft). At the beginning of September 1980 the surface site preparation and pre-grouting work was completed by JWR, and TMCI was able to commence with Stage I of the shaft sinking program - the drilling of the pilot hole.

Jan 1, 1982

By J. M. Oblak, W. A. Owczarski

A cellular appearing recrystallization product formed by annealing a cold-worked nickel-base super-alloy at 1800°F has been studied by electron nzicroscopy. Prior to deformation, an equilibrium micro-structure of fcc matrix y and cuboidal ,,&apos;, Ni (Al, Ti), precipitates of CuzAu structure had been established by an age at 1825°F. The strain-free recrystallization cells consist of very large rodular y&apos; particles in a y matrix. They precipitate is oriented and coherent both before and after recrystallization. The results showed that y&apos; coarsening accompanies recrystallization at 1800°F. However, it does so as a secondary effect and does not necessarily take place at lower temperatures. The structural similarity of this reaction to cellular precipitation in other systems indicates that lattice strain may also play a significant role during some cellular precipitation reactions. THERE have been numerous microstructural investigations of recrystallization in single-phase materials but two-phase systems have received much less attention. The second phase can either remain inert or be altered along with the matrix during recrystallization. If the second phase is an oxidelm3 or a relatively inert pre~ipitate,~, recrystallization is retarded when the interparticle spacing is less than 1 p. Prior to the onset of recrystallization, these materials show a well-polygonized substructure with the subgrain size limited by the interparticle spacing. Since recrystallization by the motion of preexisting grain boundaries6 is not observed, retardation has been related to particle pinning of the subboundaries. This pinning prevents coalescence&apos; or growth8 of subgrains to a critical size (formation of a high-angle boundary) necessary to initiate recrystallization. In a material such as a nickel-base superalloy both y matrix and y&apos; precipitate are altered by the recrystallization reaction. Haessner et al.&apos; studied the recrystallization of a cold-rolled Ni-Cr-A1 alloy by electron microscopy. The material was initially cold-rolled in the supersaturated condition. upon annealing at 750°C, immediate precipitation of 7&apos;occurred. Presence of this 7&apos; greatly retarded the onset of recrystallization which eventually took place by the development of randomly oriented, strain-free grains. The original •/ was dissolved at the recrystallization interface and reprecipitated as oriented, coherent par-tiles in the new grain. Recrystallization caused a refinement of .)&apos; particle size. Recently ~hillips&apos;&apos; investigated recrystallization of Ni-12.7 at. pct Al. Reduction by cold rolling presumably elongated the p&apos; precipitate into lamellae that remained coherent with the matrix. After recrystallization at 600" to 750°C, there was no unusual change in y&apos; particle size al- though there was a tendency toward clustering along the prior rolling direction at 750°C. Above 750°C, the recrystallized grains were generally free of precipitate. Studies in the somewhat analogous Cu-3.23 wt pct CO" and Cu-2 wt pct&apos;2 systems demonstrated that the coherent cobalt-rich fcc precipitate in these alloys obstructed softening, initiation, and completion of recrystallization. The precipitates were deformed into lam~llae during rolling and those of diameter less than 250A remained coherent. Recrystallization took place by the growth of new grains into the recovered or poly-gonized material. In the first study," both matrix and precipitate reoriented in the same manner upon passage of the recrystallization interface. There was no change in particle size or morphology. Tanner and ~ervi,~ on the other hand, observed that motion of the recrystallization fronts was strongly hindered by the pinning action of coherent precipitates in the deformed material. Particles in contact with a pinned boundary coarsened and coalesced leaving a denuded zone in the unrecrystallized region. When the number of pinning points was sufficiently reduced by coalescence, the boundary swept past these particles and through the denuded zone. The authors1&apos; considered this as a variation of discontinuous precipitation with both chemical driving force and deformation strain energy contributing to recrystallization. Preliminary observations by the present authors had revealed that recrystallization in Udimet 700, a nickel-base superalloy, occurred in an entirely different manner. Optical metallography showed that the recrystallized product formed as cellular colonies containing coarse y&apos; particles elongated in the direction of cell growth. In this investigation the structural features of this reaction were investigated by transmission electron microscopy. EXPERIMENTAL PROCEDURE As-received I$-in. rounds of Udimet 700* were (wtpct) 18.4 15.2 4.95 4.42 3.43 0.06 0.031 0.14 Bal. solution-annealed for 4 hr at 2150" and then fast air-cooled. An initial y-~&apos; structure was established by a 4-hr age at 1825°F followed by a fast air,cool. Essentially the equilibrium volume fraction of ?&apos; at 1825°F is precipitated within 4 hr. Microstructural examination showed no measurable increase in the amount of precipitate after longer aging times. Deformation consisted of swaging to 52 pct RA with 6 pct reduction per pass at room temperature. To reduce the precipitation potential to a negligible amount, recrystallization anneals were conducted at 1800"~ (982"~). Microstructures were investigated by optical and transmission electron microscopy. To prepare foils for electron microscopy, the material was first sliced into 30-mil slabs parallel to the swaging direction. Discs were dimpled and electrolytically cut from

Jan 1, 1969

One-half fare return rate again available to members of the Institute and dependent members of their families. DON&apos;T FORGET YOUR RAILROAD CERTIFICATE Over 300 members of the A. I. M. E. and dependent mem-bers of their families attending the last annual meeting se-cured return railroad tickets at half price by means of the railroad certificate, thereby saving several thousand dollars. DIRECTIONS 1. Be sure to obtain your railroad certificate when pur-chasing your railroad tickets for the meeting or you will lose money. 2. Purchase your tickets at regular one-way adult tariff fare between the dates specified for your territory in the accom-panying table. 3. Present your certificate for valida-tion at Institute headquarters in New York immediately after registration. Certificates will be validated on Feb. 21 and 22 between 9 a. m. and 5.30 p. in. 4. You must use your return half-rate tickets not later than Feb. 27. 5. Use the same route for both going and return trip. PLEASE REMEMBER THIS 1. Unless-250 certificates are validated at headquarters it will be impossible for anyone to secure the reduced rate. 2. Unless you secure a certificate you cannot take advantage. of the half-rate return fare, even if the necessary 250 certificates are validated.

Jan 2, 1928

In a letter to Prof. L. P. Breckenridge, chairman, urging the continua-tion of the Fuel Conservation Committee of the Engineering Council, Edwin Ludlow calls attention to the millions of tons of culm in the an-thracite region that can be utilized in properly equipped plants and the electricity generated transmitted to. New York, Philadelphia, and other nearby cities. He says: "Experiments have shown the economic value of No. 4 Buckwheat, which passes through a 1/16-in. mesh and over a 1/32-in. mesh. Culm passing through a 1/32-in- mesh is pulverized until it passes through a 300 mesh and is then burned under boilers. Two plants of this kind are now in successful operation in the anthracite region. Another rather interesting experiment was using 30 per cent. of this pulverized coal with 70 per cent. of oil in the generation of steam in the place of pure oil. This utilization of the fine anthracite, which has always been a waste product and thrown away, is not only fuel conservation in its truest sense but in the case of a bituminous strike, when no strike is contemplated in the anthracite field, will furnish fuel to manufacturing plants that otherwise would be obliged to shut down."

Jan 12, 1919

By Charles Clayton

A study of the literature shows that the greater part of research work on annealing of cold-worked iron has been for the purpose of studying the effect on grain-size and properties other than hardness. No reference has been found of experimental work of the same nature as that explained in this short paper. The material used throughout was Vismera iron, an iron containing 0.03 carbon and made several years ago by the Inland Steel Co. One-half inch stock was cut into cylinders ¾ in. long. The cylinders were then compressed for 60 sec. in a Reihle testing machine under loads as indicated below: COMPRESSION, AVERAGE LENGTH, SERIES POUNDS INCHES 1 10,000 0.693 2 15,000 .599 3 20,000 .501 4 25,000 .426 5 30,000 .378 6 35,000 .351 7 40,000 .311 Fig. 1 shows these cylinders after compression. Each series con¬sisted of 17 specimens lettered 0, B, C, D, E, E:, G, H, I, J, K, L, M, N, P, Q, and R. One specimen of each series was reserved for study of the properties in the cold-worked condition. The other specimens were annealed in a Hump furnace for ½ hr. at the following temperatures:

Jan 2, 1926

By C. B. Murray

"OUR experience from time immemorial has been that the furnace sampler and analyst usually find one or two, and sometimes more, points less iron than the Lake Superior shippers, but, fortunately, their contracts are `iron bound&apos;; and, generally speaking, analyses at this end are practically useless, for we are obliged to settle on basis figured out to a half dozen decimals, and, as a rule, without recourse in the event of flagrant discrepancies between the two analysts." The above is an extract from a letter received by the writer of this paper from the President of an iron company, discussing the question of sampling. It so well expresses the opinion of many blast-furnace owners and managers that it seems a suitable text for a discussion concerning the need of uniform methods of sampling. If all railroad cars were loaded the same, if all stock piles were the same shape, and if ore was received regularly each day, the problem of sampling would be comparatively simple. The conditions, however, vary widely at different places, and at the same place from day to day, and it is probably impossible to determine upon any one method of sampling which will apply at all times and in all cases. However, it would seem as though some general principles might be suggested which would tend to overcome these wide differences in samples of the same thing. The sampling of a cargo of iron ore begins at the breast of ore in the mine. The ore is sampled from the mine cars, from the stock piles, and from the railroad cars at the mines. At the lower Lake ports the ore is practically all sampled from the boat, occasionally from railroad cars. At the furnaces, the ore is nearly all sampled from railroad cars and occasionally from stock piles.

Jan 1, 1914

By W. H. Andrews

THE constantly increasing consumption of Portland cement in the State makes the above subject of partlcular interest at this time. The increasing demand is due to the rapid settling of the country and the fact that new uses are being found frequently for the above named commercial article. The industry is one that is bound to grow as the population increases, and this being the case, it may be well ,it this time to consider briefly- the matter of the chemical and physical com-position of proper raw materials. A cement mixture is one that contains approximately 75 per cent. carbonate of lime, 15 per cent. Silica, and 5 to 7 per cent. alumina, and iron oxides; the other elements being magnesium carbonate and small quantities of the rarer elements general associated with limestone deposits. It call be readily seen from the above composition that it is possible to build up this mixture from a number of different materials. First.-A rock high in carbonate of lime can be mixed with clay or shale in such proportions as to bring about the desired composition ; the chief difficulty in the building up of such a mixture is that it is hard to find a clay or shale in which the silica-alumina, ratio lies .between 4 to 1 and 3 to 1 and at the same time shows physical characteristics which will make certain its adaptability to the practical manufacture of Portland cement. Very often clays or shales are found which are right chemically but contain large quantities of grit or free silica. Such clays or shale must be ground to extreme fineness in the mixture in order to make the silica and alumina combine with the lime when the mixture is burned. On the other hand, clays or shales are often found that have the proper physical characteristics but in which the amount of alumina, is too great.

Jan 9, 1913

By R. G. Knickerbocker, R. R. Lloyd, W. T. Rawles

In July 1041, the Experiment Station of the Bureau of Mines at Boulder City, Nevada, began a study of methods of producing magnesium metal from magnesium oxide, with particular emphasis upon the direct use of magnesium oxide in a molten chloride electrolyte. The program was extended to include pilot-plant study of chemical methods for the preparation of pure magnesia from dolomites and magnesites. Preliminary work on the production of magnesia from dolomite was done with the cooperation of Kobert D. Pike and the Harbison-Walker Refractories Co. These investigatipns included a cooperative arrangement whereby the Bureau tested certain steps of the Harbison-Walker-Pike process. The process finally developed by the Bureau includes the precarbonation mixing step of the H.W.P. flowsheet prepared by Mr. Pike. Two magnesium-reduction plants using MgO as a source of magnesium were being constructed in the West, and they enormously stimulated interest in various dolomite deposits as possible sources of the magnesia needed by them. This paper describes one phase of this program—the operation of a pilot plant for producing cell-grade magnesia from dolomite from Sloan, Nevada. The process developed for this purpose is an adaptation of a well-known process first proposed by Clossonl and modified by others,2"1 involving the use of calcium chloride and carbon dioxide. It is based on the following chemical reactions: MgCl2 + Ca(OH)2 = CaCl2 + Mg(OH)2 [1] Mg(OH)4 + CaCU + CO2 = MgCl2 + CaCO, + H2O [2] The first reaction proceeds almost completely to the right because of the much greater solubility of calcium hydroxide as compared with magnesium hydroxide. Under equilibrium conditions, when only the four components shown in reaction I are involved, the ratio of magnesium to calcium in solution, according to solubility-product relationship, will be: This reaction is the basis of present commercial processes for recovering magnesium from sea water and dolomite. The second reaction is, in effect, a reversal of the first reaction produced by adding COz. This reaction also proceeds nearly to completion, the solubility relationship being: Ca where P indicates partial pressure of COz, in atmospheres.

Jan 1, 1944

By M. G. Corson

UNTIL recently, the characteristics of cast copper have seldom been studied and such study as has been made has been inadequate. The prevailing idea seems to be that since pure copper is infrequently used in the cast state, there is no reason for trying to establish precise data regard-ing the properties of the metal in that condition. On the other hand, wrought copper either in the shape of sheets or wire rarely fails in service, except as a result of the action of reducing gases at the annealing tempera-tures, so there is little incentive to study the properties of the cast shapes -cakes, bars, ingots and billets-for-the sake of tracing the causes of failures of the wrought metal to the cast stock. However, about a decade ago, wirebars of American origin often cracked badly while being hot-rolled in mills in Germany. This led to four investigations: by E. Seidl and E. Schiebold,1 by P. Siebe,2 by O. Bauer and G. Sachs,3 and by G. Welter.4 These authors came to the conclusion that failures in hot-rolling were due essentially to the presence .of a wide zone of columnar crystals between the thin layer of fine crystals near the surfaces of the bars and the large but more nearly equiaxed grains of the central core. This macrostructural defect was explained in turn by the probable upward deviation in the temperature of the metal while it was being poured into the molds of the casting machine. More recently, cast shapes of copper were investigated by N. P. Allen5 and by C. Blazey.6 Allen examined vertically cast wirebars and hori-zontally cast wedge cakes of electrolytic tough-pitch copper, but only for the distribution of densities; i.e., porosity. Blazey was interested mainly in the mechanism of the variations in electrical conductivity clue to cold-work and anneal, but gave introductively an account of the composi-tion and average densities of the tough-pitch horizontally cast wirebars used. As far as the present writer can tell, no investigation was under-taken for the sake of presenting a picture of the distribution of mechanical characteristics in the wirebars of copper, tough-pitch or otherwise.

Jan 1, 1938

By R. W. Lai, I. Iwasaki

Some physical as well as chemical aspects of the methods of dispersion of various corn starches and their derivatives are investigated and their effects as depressants in anionic silica flotation of iron ores ascertained. The adsorption behavior of starches on hematite and quartz is strongly affected by their molecular sizes as inferred from viscosity measurements. The implication of this behavior on flotation and flocculation is discussed. Mechanical treatment of causticized starch solutions is shown to reduce the starch requirements in flotation appreciably. Starches and starch products have been used as depressants for iron oxides both in amine flotation of siliceous gangue and in soap flotation of activated silica from iron ores. Cooke and his associates1,2,3,4 extensively studied the effects of certain starches and their derivatives in the amine flotation of iron ores at near-neutral pH and found that although most starches were good flocculants for hematite and de- pressed that mineral more or less selectively1,3 they did not flocculate quartz suspensions. Further, that replacement of active groups in corn starch modified its behavior as a flocculant for hematite or quartz suspensions1 and as an iron oxide depressant. Schulz and cooke2 showed that many starches and their derivatives, except aminoethyl starch, were abstracted more effectively by Brazilian hematite than by quartz and that the adsorption of a starch product was strongly dependent on pH and on the presence of calcium ion. In soap flotation of calcium-activated quartz, certain starches and their products have been used to depress iron oxides.5&apos;6 The behavior of these substances in flotation appears at times to be at variance with those previously reported and especially with results obtained in the Mines Experiment Station laboratories. This is attributed, at least in part, to differences in flotation procedure and flowsheets and in the method of preparing the depressant. Most of the investigators referred to above, solubilized the starch products by autoclaving at 120°C. Others have used acid,7 alkali8 or oxidizing agents.9 In view of the limited knowledge of the depressing action of starches and their derivatives in iron ore flotation, particularly in soap flotation of calcium-activated silica at high alkalinity, an extensive study

Jan 1, 1965