Search Documents

By Harry C. Swanson

TRANSPORTING concrete from mixer to forms has always been a problem. Twenty-five years ago this task was generally accomplished by means of wheelbarrow or concrete buggy. On large dam jobs, as the number of these projects increased, the gantry crane or highline came into use. Today several methods of handling concrete are employed on smaller surface construction jobs, for example, transit-mix trucks or dumpcrete trucks, which have crawler cranes with buckets for placing concrete into forms. In 1944, during early stages of developing Mather mine A shaft, several large underground concrete jobs were necessary. At this time the Cleveland-Cliffs Iron Co, purchased the first pump-crete machine, introduced by the Chain Belt Co. of Milwaukee. The machine was used to pour approximately 200 cu yd of concrete for a dam, or bulkhead, located 400 ft from the shaft. Concrete was mixed on surface, lowered down the shaft 1000 ft in a 2-cu yd bucket hung under one skip, spouted into the bowl of the pumpcrete machine from the bucket, and pumped directly into the forms. Since the day of the first pipeline concrete in 1944 to the present time, other equipment and other methods have been developed to permit transportation of concrete by pipeline through vertical and horizontal distances totaling 1 mile from mixer to forms. Much of the efficiency in present handling of underground concrete can be credited to the Bethlehem Cornwall mines, where concrete was transported through 6-in. pipe for great distances down an inclined shaft and along levels into forms.' During initial development of Mather mine B shaft, with concrete work under way on two or more levels at one time, the pneumatic concrete placer, Fig. 1, was selected as best adapted for underground concrete transportation. The 3/4-cu yd pneumatic placer is a small machine readily moved from one location in the mine to another. It can be equipped with two sets of mine car wheels, which will permit moving on regular mine tracks. It is therefore possible to send concrete through the pipe at great velocity; the pipeline is clean after each shot except for the film of cement adhering to the inside. With the proper slump in the concrete, it is possible to shoot concrete 2000 ft with this machine, using the mine supply of compressed air at 95 psi. This equipment was first used at Mather mine B shaft to concrete slusher drifts, Figs. 2 and 3, and finger raises located about 2000 ft from the shaft. In several instances there were bends into crosscuts and up vertical distances into the forms. For the first pours two placers were used. The first was located near the shaft where the concrete could be spouted into it from a 2-cu yd concrete bucket on the cage. The second was set on the side of the drift at a point approximately 1500 ft from the shaft. The concrete was shot directly into the second placer from the first unit and from the second machine directly into the forms. After completion of several pours with the two machines, a trial pour with only one placer located at the shaft proved that the second placer could be eliminated. Since then all pours have been successfully completed with only one placer underground. As production of iron ore from the mine increased and the development program expanded, use of the cage for handling mine supplies and concrete became a major problem. This brought about the first attempt at shooting concrete vertically down the shaft for 2600 ft. A 6-in. pipeline with victaulic couplings installed during shaft sinking was used for the trial. One placer was set on surface 250 ft from the collar of the shaft so concrete could be spouted directly into it from the mixer. This machine shot the concrete 250 ft horizontally on surface to the shaft, 2600 ft vertically down the shaft, and 100 ft horizontally into the second placer located near the rib of the shaft station or plat. The second machine shot the batch into the forms, about 2000 ft. Total distance horizontally and vertically was 4800 ft. The entire time cycle for a ¾-cu yd batch of concrete from the mixer on surface to the forms underground totaled about 5 min. During the past two years the two-placer method from the mixer on surface to the forms underground has proved a very efficient means of transporting underground concrete. Advantages of using pipeline concrete are as follows: 1—Interference with normal mining operation is eliminated. When the cage, skips, mine cars, or mine openings are used for transporting concrete and materials used for making concrete, mine operation suffers in one way or another.

Jan 1, 1955

By AIME AIME

Air transport supplants the old methods. The 3-piece canoe fits in the plane and likewise makes possible not a bad division of labor over a 5-mile portage

Jan 1, 1933

By P. R. Sperry, A. P. Beck

It has been shown1 that in poly-crystalline strips of high purity aluminum with a fairly random orientation distribution, grain growth progresses gradually until the average grain diameter reaches a value approximately equal to the strip thickness. Recent work at this laboratory led to the realization that grain growth might be impeded to a considerable extent in the presence of a sharply defined texture, where orientation differences between neighboring grains are small. In order to investigate this effect the following experiment was carried out with the same lot of high purity aluminum previously used for grain growth studies in randomly oriented material.' Very large grain size was developed by grain growth at 650°C in specimens of 0.200 in. thickness. These specimens were then rolled to a thickness of 0.050 in. or 1.25 mm—a reduction of 75 pct. In the rolled strip each large grain corresponded to an elongated area easily identified by etching. After annealing for 1 to 25 min at 600°C and re-etching, these elongated areas were again recognizable. Within each area, corresponding to a single large grain before annealing, there formed by recrystallization a multitude of new grains with a fairly well developed preferred orientation. The orientation and the size of the new grains formed in areas corresponding to different large grains, varied widely depending on the orientation of the parent grains with respect to the rolling direction and the plane of rolling. Many areas were found where the average grain size was considerably smaller than the specimen thickness. Such an area occurred in a specimen cut in half before annealing. One half, containing a portion of the area in question, was annealed 1 min at 600°C, the other half, with the remaining portion of this area, for 25 min at the same tempera-Aluminum killed low carbon steel, § which is now used extensively for severe deep drawing or other difficult forming operations, is unusual in that its grain structure, after cold reduction and box annealing in accordance with conventional continuous sheet or strip mill practice, often is elongated, although at times it is equiaxed. Since this unusual structure has been found superior for many, but not all, severe forming operations, recrystallization of the steel, both at constant temperature and on continuous heating, was investigated and compared with that of rimmed steel in the hope that something might be learned about the mechanism of, and the factors controlling, the formation of such elongated grains. In this structure, the grains are elongated both in the lengthwise direction of the strip and transverse to this direction, even though nearly all of the extension in both hot and cold rolling is in the lengthwise direction. The grains are thus roughly pancake-shaped, being longer and wider than they are thick, as observed also by Burns and McCabe,1 and as illustrated by the typical structures shown in Fig 1. Fig la, representing a conventional longitudinal section, shows the length and thickness of the grains, whereas Fig Ib shows their length and width as seen by examining a section parallel to the sheet surface. Both illustrate the very irregular grain boundaries usually associated with the elongated grain shape. A finer equiaxed grain structure in this same grade is shown in Fig Ic. Either the elongated or the equiaxed structure may be present in the annealed product, and in rare instances the two types may coexist in a single specimen, as shown in Fig 1 d. Isothermal Recrystalliza-tion of Rimmed and Alamimum Killed Steel An aluminum killed steel known to have an elongated grain structure after conventional processing (Steel B, Table l), was selected for the initial recrystallization studies; for comparison, a rimmed steel, A in Table 1, was used. Samples of each in the form of hot rolled strip 0.075 and 0.095 in. thick, respectively, were cold rolled on a small laboratory mill in steps of about 0.010 in. per pass to obtain total reductions of 40 and 60 pct. Small pieces of the cold reduced strip were heated in lead at selected constant temperatures for one of several periods of time, then cooled in air. Rate of heating in the lead was, of course, very fast. Hardness of the cooled specimen was measured and a longitudinal section examined metallographically. Isothermal recrystallization curves for these two steels at 1050°F, based on hardness of the air cooled specimens, are shown in Fig 2 in which the amount of recrystallization corresponding to each plotted point is indicated. The marked difference in the behavior of these two types of steel is evident. After a corresponding amount of cold reduction, the rimmed steel recrys-tallizes in a much shorter time than the killed steel and the shape of its recrystallization curve, (plotted on a logarithmic time scale), is very different. The curve for rimmed steel indicates that recrystallization is analogous to isothermal transformation of aus-i.enite in that it proceeds at a progressively faster rate up to some 50 pct recrystallization, then at an increasingly slower rate. For the aluminum killed steel, however, the start of

Jan 1, 1950

Jan 1, 1930

By Jaakko Kajamaa

Neutron diffraction was applied to determine sheet textures by the transmission method. Cold-rolled and recrystallized copper sheets were investigated. The amount of cube texture was determined for three compositions, in which the phosphorus content was, respectively, 0, 0.005, and 0.03 wt pct. The heat treatment was in every case 8 sec at 650°C. In the two latter cases the cube texture was prevented. In addition a comparison with the X-ray diffraction transmission method was made with the 96 pct cold-rolled copper sheet. Outer parts of both (111) pole figures can be considered to be rather identical. This is seen from the fact that the intensity ratio ITD/120" was 0.45 for neutron diffraction and 0.40 for X-ray diffraction. Differences between the methods were discussed in detail. Features peculiar to neutron and X-ray diffraction in texture studies were listed and compared. In this work neutron diffraction was applied to determine sheet textures. Specifically, it was desired to ascertain whether this method can be used to reveal differences when compared to other methods. In addition, the amount of the cube texture in copper sheets was determined as a function of phosphorus content. Previous applications of neutron diffraction to texture problems include the following: nickel wires,' wire of some bcc metals,' and uranium bars.3 In the neutron diffraction technique the greatest difference is in the sample—its method of production and its volume. A sample needs no treatment and its volume is roughly 105 times larger than the volume of an X-ray diffraction sample. The cold-rolled sheet was investigated both by neutron diffraction and by X-ray diffraction, because it is expected that, due to large number of defects, possible differences in the results of the two methods would be revealed. It is a well-known fact that X-ray lines show broadening when cold-worked. Analysis has shown that this is based chiefly on small crystalline size, micro-stresses, and/or faults.4'5 Neutrons are sensitive to the above-mentioned disturbing factors as well, but circumstances in diffraction are different from the X-ray case. Because the sample represents a larger volume, the result is an average over that volume. In addition, it can be assumed that the sample has preserved its original structure, because it needs no special preparation. The particular limitation of neutrons is the relatively low neutron intensity available from nuclear reactors. This decreases the resolution as compared to the X-ray diffraction methods. Furthermore, absorption mainly reduces diffracted X-ray intensity, while multiple scattering effects, i.e., secondary extinction, disturb neutron diffraction. SO neutrons and X-rays behave in a different way when interacting with matter. As in other structural investigations, one can utilize this difference in texture studies as well. One cold-rolled and three recrystallization textures in copper sheets were investigated by neutron diffraction. The samples were produced at the Outokumpu copper factory to the specifications shown in Table I. The paper is divided into five parts. The first deals with the theory of the measurement. In the second, experimental procedures are described. Results are presented in the third part. Both cold-rolled and re-crystallized samples are studied. Discussion is in the fourth part, and finally in the fifth part some conclusions are drawn. 1) THEORETICAL CONSIDERATIONS Properties peculiar to neutron diffraction are the following: a) the scattering length varies greatly between one element and another; b) many of the elements do not absorb neutrons appreciably. In this connection it is of primary interest to know the interaction of neutrons with lattice imperfections. As with X-rays this problem leads to diffraction analysis of deformed and recrystallized metals. From the physical point of view the main difference is that neutrons are scattered by nuclei (magnetic scattering is not considered here), whereas X-rays are scattered by electrons. The features peculiar to neutron and X-ray diffraction methods in texture studies are listed in Table 11. Pole figures are an important tool in performing structural analysis of deformed or recrystallized metal. Present texture research technology requires pole figures which are as precise as possible. The choice between these two methods depends on the technical information which is required. The X-ray diffraction transmission technique may give results which are not necessarily representative of the average physical state of the sample. Although foil samples normally contain enough crystallites for diffraction, they may not necessarily represent the whole structure. An example of this problem is the frequently observed difference between the "surface" and the "inside" texture of a sample. The production of foil samples may disturb the original structure of the parent material. The selection and orientation of the foil from the sample is quite arbitrary. Normally, a highly deformed piece of metal has several texture components. Different components are deformed in a slightly different manner. This is a re-

Jan 1, 1969

By J. E. Worthington, W. H. Spence, I. T. Kiff

Gold was discovered at the Haile mine in Lancaster County, South Carolina, in 1827 or 1828, and since that time the mine has been worked intermittently by both open-pit and underground methods until its forced closure in 1942 by World War II. Production figures are incomplete, especially for the early years, but the total gold produced is estimated to have been greater than 200,000 oz. Thus, the Haile mine has been the most productive gold mine in the eastern United States. The upper, residually enriched ores were relatively rich, but the bulk of the production has come from the mining of lower grade ores. General Geology The Haile mine is located in late Precambrian or early Paleozoic rocks of the Carolina slate belt at the edge of the Atlantic Coastal Plain [(Fig. 1)]. The metamorphic grade is lower greenschist facies and the rocks have been folded into a sequence of northeast-trending isoclinal folds. The gold is associated with siliceous, pyritic, and kaolinized felsic pyroclastic and tuffaceous rocks in an interbedded volcanic and volcanoclastic sequence of felsic to mafic tuffaceous rocks and argillaceous sediments [(Fig. 2)]. The ore bodies occur in two northeast trending zones approximately 500 m apart; each zone is 30-70 m wide and 600 m or more in length, with possible extensions to the east beneath the Coastal Plain sediments. Mineralogy. Gold in the Haile mine is always associated with siliceous and/or pyritic ores. The gold occurs in at least three states: As native gold as originally deposited; as residual gold derived from the breakdown of pyrite; and as gold included in pyrite. Major associated minerals in addition to quartz and pyrite are kaolinite, sericite, and iron oxides. Minor molybdenite, arsenopyrite, pyrrhotite, copper sulfides, sphalerite, rutile, and topaz are also present. Petrology. The gold-bearing ore zones vary from highly siliceous rocks to pyritic massive sulfide lenses. This variation is most easily seen today along strike from the Haile pit to the Red Hill pit. Ore grade material still exposed in the wall of the Haile pit consists of a highly siliceous and very thinly bedded rock containing minor pyrite. Along strike, the character of the mineralization changes to pyritic massive sulfide lenses occurring interbedded with siliceous horizons at the Red Hill pit. The siliceous rocks vary from the thinly-bedded material as just described from the Haile pit to silicified fragmental-appearing rocks to totally recrystallized cherty rocks lacking any recognizable primary features. Scattered, apparently at random, throughout the very thinly-bedded and very fine-grained ore face of the Haile pit are seemingly anomalous silica-rich clasts or concretions up to 5 cm in diameter which will be discussed later in this paper. Alteration. One of the most striking features of the Haile deposit is the alteration mineral assemblage which is intimately associated with the siliceous and pyritic ores. This altered material has been intersected in drill core at depths greatly exceeding the modern weathering profile and is, therefore, of hydrothermal origin rather than from supergene processes. This "sericite," actually a fine-grained mixture of sericite, kaolinite, and quartz, can be shown to stratigraphically underlie the gold- quartz-pyrite zone, and is well exposed in the open pit just southeast of the Haile and Bumalo pits. Relict textures indicate that this highly altered material was originally a felsic ash flow. Other similar alteration zones have been found in outcrop and drill core underlying the remaining ore bodies. Thus each of the mineralized zones consists of two parts: A siliceous and/or pyritic gold-bearing ore zone which is stratigraphically underlain by a zone of high alumina minerals, in this case sericite and kaolinite along with variable amounts of quartz. A green chrome mica, presumably fuchsite, is present in trace amounts in the high alumina zone. Genesis An adequate model to explain the origin and distribution of the gold deposits in the Carolina slate belt is presently lacking. Worthington and Kiff1 suggested a volcanogenic origin for certain gold deposits in the North Carolina slate belt from the waning exhalations of felsic volcanic piles. They also pointed out that such an origin has similarities to many epithermal precious metal deposits located in more recent volcanic piles in the western United States. A further key to the understanding of the genesis of the gold mineralization at the Haile mine is the close association of the mineralization in siliceous and sulfidic horizons to the genetically related and stratigraphically underlying high alumina alteration. Such high-alumina alteration is common around felsic volcanic centers in the Carolina slate belt and the mineralogy as seen today consists of some combination of kaolinite, sericite, pyrophyllite, kyanite, andalusite or sillimanite depending on the local prevailing grade of metamorphism. Accompanying the high-alumina alteration are large quantities of pyrite and iron-oxide minerals as well as characteristic minor accessory minerals often including base metal sulfides, fluorine-bearing minerals (topaz, fluorite, apatite), titanium-bearing minerals (ilmenite, rutile),

Jan 1, 1981

By Carrol M. Beeson

Reasons are given for using a Boyle's-law porosimeter in conducting core analysis for either routine or research purposes. Among other things, it is pointed out that such a porosimeter permits the measurement of all basic properties on the same sample, thereby eliminating the sources of error inherent in the use of adjacent samples. References are made to investigations of gas adsorption on various porous materials, to show that the use of helium in Boyle's-law porosimeters reduces to negligible proportions the error due to the adsorption or desorption of the operating gas. Two Boyle's-law instruments are described. which permit accurate and rapid measurements of porosity. Schematic sketches and explanation:; are included, along with derivations of equations required in performing precise determinations. Summaries of data obtained during calibration are tabulated and analyses of the data are resented as indications of the precision and accuracy of each device. Comparisons are also shown for measurements made with each of the instruments on the same test pieces and cores. INTRODUCTION An accurate porosimeter, operating on the principle of Boyle's law. is of considerable value in the analysis of cores for either routine or research purposes. This is due primarily to the fact that the measurement of porosity with such an instrument leaves the sample free of contamination by any liquid. When used in conjunction with an extraction apparatus' for determining oil and water saturations, a Boyle's-law porosimeter permits the measurement of all basic properties on the same sample. This eliminates the sources of error inherent in the use of adjacent samples, or the necessity of determining porosity after all other properties have been obtained. Large errors may result from combining measurements made on adjacent samples in order to obtain a single property. This type of error is definitely involved when oil and water are measured with one sample, and the pore vo1ume is measured with an adjacent one. Furthermore, the source of error would be present to some extent, even if the analyst could choose the samples so they were truly adjacent from a geological standpoint. The use of adjacent samples in routine core analysis also necessarily decreases the probability of correlating core properties. For example, the chance of correlating the "irreducible" interstitial-water saturation with permeability, is bound to be greatly reduced by measuring these properties on "adjacent" samples. For research purposes, amplification is scarcely required concerning the greater flexibility of a method for measuring porosity which leaves the core free of contamination by any liquid. Even under those circumstances which require that the core be saturated with a liquid, a previous measurement of porosity with a gas is useful in determining the degree of saturation that has been attained in the process. Furthermore, for comparable accuracy, porosity usually may be determined more rapidly with a gas than with a liquid. This advantage of the Boyle's-law instrument is most outstanding when the determination time is compared with that required in obtaining porosity by evacuation of the core followed by saturation with a liquid of known density. Several porosimeters which operate on the principle of Boyle's law have been described2,3,4,5,6,7 in the literature. No comparison will be attempted between those instruments and the ones described herein. Before helium gas became readily available, Boyle's-law porosimeters were subject to an appreciable error due to the adsorption of the operating gas on the surface of the core solids. There is considerable direct and indirect evidence in the literature to support the contention that the adsorption of helium on porous solids is negligible at room temperature. In discussing the use of Boyle's-law porosimeters, Washburn and Bunting2 stated that "for most ceramic bodies dry air is a satisfactory gas, but hydrogen will be required in some instances. Helium could, of course, be employed for all types of porous materials at room temperatures or above." Howard and Hulett8 obtained evidence that the adsorption of helium was negligible at room temperatures, even on activated carbon ; for the density measured with this gas was unaffected by changes in pressure or by changes in temperature from 25 °C to 75 °C. For oil-well cores, Taliaferro, Johnson, and Dewees" obtained lower porosities with helium than with air, but apparently did not study helium adsorption. From the work of these investigators, it follows that the use of helium in Boyle's-law porosimeters reduces the error due to gas adsorption to negligible proportions. This makes it possible to construct instruments which permit the determination of porosity with (1) a high degree of accuracy, (2) with great rapidity, and (3) without contamination. THE KOBE POROSIMETER The fundamental design of the Kobe Porosimeter was developed by Kobe, Inc., which firm built about 12 of the instruments during 1936 and 1937. Since that time, seven or eight more have been constructed with their permission, making a total of about 20 that have been put into operation.

Jan 1, 1950

By E. S. Bartlett, D. N. Williams

QUANTITATIVE values for the oxidation rate of unalloyed molybdenum in air at temperatures above the melting point (1460°F) of the characteristic oxide are contained in the literature as a result of previous investigations. Lustman' reported values corresponding to 0.36 in. penetration per day (IPD) at 1500 and 1600°F in still air, noting essentially no variation in rate with temperature. Jones, Spretnak, and Speiser' reported values corresponding to 0.14 and 0.13 IPD at 1500 and 1800°F, respectively, in still air, attributing the decreased oxidation rate at higher temperatures to a lesser accumulation of the corrosive molten oxide on the surface at the higher temperature as a result of increased volatilization rate. Harwood3 ecently summarized work in the field, presenting generalized data corresponding to 0.48 to 0.96 IPD at 1800°F and 0.55 to 0.83 IPD at 1700°F in slowly flowing air. In a recent program at Battelle, it became desirable to know more about the characteristic oxidation behavior of molybdenum under varying conditions of temperature and atmosphere. Using oxidation-test apparatus designed for dynamic, continuous recording of weight change during testing,' values for the oxidation rate of molybdenum were obtained at temperatures from 1400 to 2150°F. In addition the effect of air flow on the oxidation rate was studied briefly at temperatures of 1600, 1800, and 2000°F. Exhaust of the contaminated atmosphere from the oxidation chamber was effected by an impeller pump attached to a 3/16-in.-diam opening in the oxidation chamber. The volumetric exhaust rate (cubic feet per hour) was normally maintained slightly in excess of the input rate to avoid condensation of MOO,,' on the sample suspension rod. The entering atmosphere was preheated prior to admission to the oxidation chamber by a 1 1/2-in.-diam cup packed with shredded asbestos. The experimental data are presented in Table I. Comparing conditions 2 and 3 (taking into account the temperature difference) and conditions 8 and 9. shows that in the absence of forced exhaust an atmosphere of moving air results in greater oxidation rate than a stagnant atmosphere. The use of forced exhaust, as shown by comparing conditions 3 and 4 and conditions 14 and 15: resulted in an even greater increase in oxidation rate. By virtue of the size of the atmosphere input and exhaust openings, it was calculated that the exhaust velocity was about 60 times that of the input velocity for essentially equal volumetric flow rates. Because of the proximity (about 3/4 in.) of the exhaust port to the specimen, it is logical to assume that cleansing of the atmosphere immediately surrounding the specimen was accomplished much more efficiently by the exhaust flow than by the input flow at a constant-volumetric flow rate. Also, it can be seen by comparing conditions 4 through 7 and conditions 11 through 13 that increasing the rate of atmosphere flow (by increasing input velocity with a proportional increase in exhaust velocity) above some optimum value has little, if any, further effect on the oxidation rate. These results suggest that there is a maximum oxidation rate for molydenum at a given temperature which is obtained when conditions are maintained such that the partial pressure of MOO3 in the atmosphere surrounding the specimen is at a low value. By controlling the partial pressure of MOO, surrounding the sample, it is possible to control the rate of volatilization of MOO:, from the surface. This, in turn, affects the rate of oxidation, since the thickness of the MOO3 layer determines the amount of oxygen which will be able to reach the reaction surface.' When the liquid oxide layer is less than some critical thickness, i.e., when the volatilization rate is high, enough oxygen is transported to the active surface to permit oxidation to proceed at the maximum rate allowed by the kinetics of the oxidation reaction. However, if the volatilization of MOO,, is suppressed, the thickness of the layer of MOO3 on the surface increases, and the diffusion of oxygen through the oxide layer becomes the rate-controlling step in the oxidation process. The lack of agreement between the present results and those of previous investigators is presumed to be due to differences in removal of the oxidation product (Moo,) from the immediate vicinity of the sample. By comparing conditions 1, 5, 10, 12 and 14, it is seen that when forced exhaust was used the oxidation rate of molybdenum increased with increasing temperature. A rapid increase was observed between 1400 and 1600°F, attributable to the effects

Jan 1, 1959

By M. R. J. Wyllie, F. Morgan, P. F. Fulton

The importance of formation factor, F, not only in electric logging but as a fundamental rock parameter has recently been stressed.',: The desirability of investigating the range of variation of the resistivity index exponent, n, in the relationship I = S ;", where I is the resistivity index and Sw the water saturation as a fraction of the void volume of a porous medium, has also been urged.3 The magnitude and variation of n with saturation and rock texture is a subject not only of theoretical interest but also one of prime importance in the interpretation of electric logs. A simple technique has recently been developed which enables both F and u to he measured with high accuracy and which may also find acceptance as a convenient method for the determination of irreducible saturation attainment in the restored state method of core analysis. Experience has taught that reproducible measurements of F are possible only if the resistance measuring electrodes are so arranged with respect to a plane face on a porous medium that they are able to make electrical contact with substantially all entry pores in that plane. In practice this may be achieved by using a platinized-platinum gauze electrode backed by some absorbent material (such as felt) which has been saturated with a fluid identical with that used to saturate the porous medium. Applicatiorl of pressure to the electrode and absorbent material then forces the gauze against the plane face of the porous medium and simultaneously squeezes saline solution through the meshes of the gauze. By this means the electrode is in continuous aqueous contact with all pores and satisfactory and reproducible low resistance contact with the porous medium is achieved. Clearly this method, although satisfactory for measurements of F, cannot be applied to the making of continuous resistance measurements on a porous medium while capillary pressure desaturation is being carried out. However, accepting the principle that for satisfactory results electrical contact must be made between a measuring electrode and all pores adja- cent to that electrude, methods of bringing electrodes into intimate contact with the surfaces of porous media were investigated. Two methods were ultimately found to be satisfactory: in the one, the metal electrode is formed on the required portion of the porous medium by the use of a metal spray gun, while in the second the electrode is painted on with an ordinary camel's hair brush. The first method has the advantage of permitting the use of any metal which can be sprayed, but has the disadvantage of requiring rather elaborate and expensive equipment. The second method is presently limited to silver electrodes although in principle other metals, e.g. platinum or gold, could be used. Moreover, the method is so simple and cheap, and has been found to be so satisfactory that it will be described in some detail. The core being investigated is cut into a right circular cylinder and is extracted and dried in the usual manner. The ends are then lightly painted with silver conducting paint* of the type used in printed electrical circuits. The quantity of paint used is not critical but the recommended, minimum compatible with entirely coating the core ends is recommended, particularly on the end that contacts the barrier. The core is then dried at atmospheric temperature for one hour or for shorter periods at any suitable elevated temperature up to about 110°C. It will be found that silver coatings so prepared are not only strongly adherent but also permeable and the core may be the core may be desaturated by the ordinary capillary pressure technique through one of the coated faces. The same permeability is characteristic also of thin metal coatings formed using the spray-gun technique. An ordinary Lucite capillary pressure desaturation cell has been adapted to a form suitable for measuring the resistivity of the saturated silver faced cores both at 100 per cent saturation (i.e., F) and at intermediate saturations down to the irreducible minimum. This has been achieved as follows: A Coors porcelain barrier, having a displacement pressure of c 30 psi was grooved across a diameter. Dimensions of this groove were c 1 mm deep and 1 mm wide at the top. The groove was then painted thickly with silver conducting paint, the paint in the groove being extended lightly over the edges of the groove for a distance of c 1 mm on each side. A 30 gauge silver wire was then arranged in the groove in a form of a spring bow, each end of the silver being held at diamet~ically opposite ends of the groove by means of plastic cement. The arc of the bow at its highest point was arranged to be a millimeter or so above the face of the barrier, while one end of the bow wire was led by means of a pressure-tight connection through the wall of the capillary pressure cell. The groove in the barrier was then Surrounded by suitably cut portions of Kleenex in the conventional manner so as to ensure capillary continuity from core to barrier, and the core placed on the barrier. The weight of the core distorted the silver spring bow and good electrical contact was thereby made between the outside of the cell and the lower painted silver electrode. Electrical connection to tile top painted silver

Jan 1, 1951

By A. S. Malicsi, I. Iwasaki

The removal of hydrophobic coatings of flotation collectors from iron ores becomes of interest when a duplex flotation process is considered for upgrading, when a pelletizing process is considered for a concentrate floated with a fatty acid or a soap collector, or when a disposal of froth products from cationic silica flotation is of environmental concern. Ozone can oxidize organic compounds rapidly, thereby removing the hydrophobic coatings of flotation collectors. Ozone is widely used for treating and purifying drinking water, waste water treatment, and for chemicals processing (Murphy and Orr, 1975; Rice et al., 1980). Its uses in metallurgical operations, however, are very sparse (Allegrini et al., 1970; Chernobrov and Rozinoyer, 1975; Ishii et al., 1970; Iwasaki and Malicsi, 1985; Matsubara et al., 1978). Yet, its high reactivity and the absence of potentially hazardous byproducts become of interest in destroying flotation reagents adsorbed on mineral surfaces or remaining in mill water for recycle or for discharge. Duplex Flotation A duplex flotation process, as applied to oxidized iron ores, would involve a fatty acid flotation of iron minerals followed by an amine flotation of the siliceous gangue from the rougher iron concentrate. Such a process has been used in the Florida phosphate fields. Fatty acid coatings cannot be removed as readily with a simple acid or alkali treatment from iron oxide surfaces as from Florida phosphates. A combination of reagents, such as lime and quebracho, lime and alkali phosphate, or sulfuric acid and oxalic acid, has therefore been proposed. In a previous article (Iwasaki et al., 1967) , the use of activated carbon was found to be effective in removing fatty acid coatings both in the duplex flotation and the pelletizing processes. The use of ozone offers another approach to the removal of fatty acid coatings from iron oxide surfaces. To investigate the possible application of the duplex flotation process, a specularite ore from Michigan analyzing 36.5% iron was used. A 600-g (1.3-1b) sample was ground in a laboratory rod mill together with 250 g/t (0.5 lb per st) of sodium silicate to -150 µm (-100 mesh). This was transferred to a Fagergren laboratory flotation cell, and deslimed four times at 20 µm (quartz equivalent). The deslimed pulp was transferred to a laboratory conditioner, diluted to 40% solids, and conditioned with 250 g/t (0.5 lb per st) of soda ash and 250 g/t (0.5 lb per st) of oleic acid. The conditioned pulp was then transferred back to the Fagergren cell, floated until barren of froth, and the rougher froth product was returned to the cell and cleaned. The results are presented in Table 1. The cleaner concentrate at this point analyzed 45.3% Fe. The cleaner concentrate coated with fatty acid was transferred to a 2-L (0.53-gal) beaker. While the pulp was agitated with a glass T-stirrer, ozone was bubbled into the agitated pulp for 15 minutes at a rate of 10 mg/min (0.00035 oz per min) ozone (250 g/t or 0.5 lb per st 03 feed). It was observed that the pulp ceased to froth after about 10 minutes. The amine flotation of siliceous gangue from the ozonated pulp was carried out first by conditioning with a dextrin, a commonly used starch depressant for iron oxides. This was followed by flotation with a stage addition of an ether amine at increments of 100 g/t (0.2 lb per st). Three stages were required to float the siliceous gangue to near completion. The three froth products were combined and cleaned twice. When the cationic flotation Rougher, Cleaner 1 and Cleaner 2 cell products were combined, an iron concentrate analyzing 64.5% iron was obtained at an overall iron recovery of 72.8%. Pelletizing Fatty acid flotation concentrates have been pelletized successfully in northern Michigan mills. But at other locations, fatty acid coatings on iron flotation concentrates proved so undesirable in agglomeration that other methods of concentration had to be sought. For example, a sinter mix containing iron ore concentrates upgraded by fatty acid flotation resulted in decreased productivity. This occurred because the micropellets of particles with the hydrophobic coating are less tolerant of moisture. Thus, the bed permeability is lost (Beebe, 1965). The agglomeration of concentrates obtained by the fatty acid flotation alone, and the hydrophobic coatings destroyed by ozonation or by the duplex flotation process, is not expected to cause any difficulty since the surfaces of the concentrates would be hydrophilic. Removal of the fatty acid coating with activated carbon, indicated by the loss of floatability, was shown to restore the decrepitation temperature of wet balls during drying cycle (Iwasaki et al., 1967). Disposal of Cationic Silica Flotation Froths Recent demands of iron blast furnaces place the silica content of the magnetic taconite pellets at about 5%. Conventional process for magnetic taconite involving fine grinding and magnetic separation often produces magnetic concentrates analyzing in excess of 5% silica. This is due to the presence of the middling grains of siliceous gangue and magnetite. Cationic silica flotation of magnetic taconite concentrates (DeVaney, 1949) may be used to reduce the silica content. But the amine coating on siliceous gangue becomes of environmental concern when the flotation tailings are discarded in tailing ponds.

Jan 1, 1986

By M. R. J. Wyllie, F. Morgan, P. F. Fulton

The importance of formation factor, F, not only in electric logging but as a fundamental rock parameter has recently been stressed.',: The desirability of investigating the range of variation of the resistivity index exponent, n, in the relationship I = S ;", where I is the resistivity index and Sw the water saturation as a fraction of the void volume of a porous medium, has also been urged.3 The magnitude and variation of n with saturation and rock texture is a subject not only of theoretical interest but also one of prime importance in the interpretation of electric logs. A simple technique has recently been developed which enables both F and u to he measured with high accuracy and which may also find acceptance as a convenient method for the determination of irreducible saturation attainment in the restored state method of core analysis. Experience has taught that reproducible measurements of F are possible only if the resistance measuring electrodes are so arranged with respect to a plane face on a porous medium that they are able to make electrical contact with substantially all entry pores in that plane. In practice this may be achieved by using a platinized-platinum gauze electrode backed by some absorbent material (such as felt) which has been saturated with a fluid identical with that used to saturate the porous medium. Applicatiorl of pressure to the electrode and absorbent material then forces the gauze against the plane face of the porous medium and simultaneously squeezes saline solution through the meshes of the gauze. By this means the electrode is in continuous aqueous contact with all pores and satisfactory and reproducible low resistance contact with the porous medium is achieved. Clearly this method, although satisfactory for measurements of F, cannot be applied to the making of continuous resistance measurements on a porous medium while capillary pressure desaturation is being carried out. However, accepting the principle that for satisfactory results electrical contact must be made between a measuring electrode and all pores adja- cent to that electrude, methods of bringing electrodes into intimate contact with the surfaces of porous media were investigated. Two methods were ultimately found to be satisfactory: in the one, the metal electrode is formed on the required portion of the porous medium by the use of a metal spray gun, while in the second the electrode is painted on with an ordinary camel's hair brush. The first method has the advantage of permitting the use of any metal which can be sprayed, but has the disadvantage of requiring rather elaborate and expensive equipment. The second method is presently limited to silver electrodes although in principle other metals, e.g. platinum or gold, could be used. Moreover, the method is so simple and cheap, and has been found to be so satisfactory that it will be described in some detail. The core being investigated is cut into a right circular cylinder and is extracted and dried in the usual manner. The ends are then lightly painted with silver conducting paint* of the type used in printed electrical circuits. The quantity of paint used is not critical but the recommended, minimum compatible with entirely coating the core ends is recommended, particularly on the end that contacts the barrier. The core is then dried at atmospheric temperature for one hour or for shorter periods at any suitable elevated temperature up to about 110°C. It will be found that silver coatings so prepared are not only strongly adherent but also permeable and the core may be the core may be desaturated by the ordinary capillary pressure technique through one of the coated faces. The same permeability is characteristic also of thin metal coatings formed using the spray-gun technique. An ordinary Lucite capillary pressure desaturation cell has been adapted to a form suitable for measuring the resistivity of the saturated silver faced cores both at 100 per cent saturation (i.e., F) and at intermediate saturations down to the irreducible minimum. This has been achieved as follows: A Coors porcelain barrier, having a displacement pressure of c 30 psi was grooved across a diameter. Dimensions of this groove were c 1 mm deep and 1 mm wide at the top. The groove was then painted thickly with silver conducting paint, the paint in the groove being extended lightly over the edges of the groove for a distance of c 1 mm on each side. A 30 gauge silver wire was then arranged in the groove in a form of a spring bow, each end of the silver being held at diamet~ically opposite ends of the groove by means of plastic cement. The arc of the bow at its highest point was arranged to be a millimeter or so above the face of the barrier, while one end of the bow wire was led by means of a pressure-tight connection through the wall of the capillary pressure cell. The groove in the barrier was then Surrounded by suitably cut portions of Kleenex in the conventional manner so as to ensure capillary continuity from core to barrier, and the core placed on the barrier. The weight of the core distorted the silver spring bow and good electrical contact was thereby made between the outside of the cell and the lower painted silver electrode. Electrical connection to tile top painted silver

Jan 1, 1951

By A. Totani, S. Nakajima, H. Okazaki

The phase diagram for the SnTe-SnSe system has been studied in the temperature range from 300° to 900°C by differential thermal and quenching techniques. The X-ray measurements were made on quenched specimens. High-temperature diffraction was also made to study the phase transition in SnSe. The system is proved to be of a eutectic type in which no intermetallic compound exists. The eutectic point is at the composition SnTeo.55 Seo.45. the eutectic temperature being 755°C. Solid solubility limits are SnTeo.6Seo.r and SnT eo. 3s Seo.6s at the eutectic temperature, and change almost linearly to SnTeo.aaSeo.lz and SnTeo.18 Seo.az as temperature decreases to 300°C. It was shown that the SnSe phase has a phase transition of the second order at about 540°C and that the transition temperature decreases with increase of the SnTe content. THERMOELECTRIC properties of tin telluride (SnTe) and tin selenide (SnSe) have been studied extensively in recent years. The variation of physical properties with composition could be of interest if these compounds form an appreciable crystalline solution. The purpose of present investigation is to confirm the formation of crystalline solution or intermetallic compound, if any, and to establish the phase diagram for this system. The crystal structure of SnTe is NaCl type with a cubic unit cell1 (a = 6.313A). The crystal of SnSe having an orthorh2mbic unit cellz (a = 11.496, b = 4.1510, and c = 4.4437A) is isomorphous with tin sulfide (SnS) which has a distorted sodium chloride structure. It has been known that SnSe has a phase at at 540°C; the transition has been assumed to be of the second order. As far as we know, only two studies on the SnTe-SnSe pseudobinary system have been reported. The conclusion obtained in these papers is that, in the composition regions near SnTe and SnSe, the system forms a crystalline solution of the SnTe structure and the SnSe structure, respectively, and that, in the intermediate region, both phases coexist. However, neither the variation of the solid solubility vs the temperature nor the liquidus and solidus were investigated. Hence present writers have attempted to determine the phase diagram of the system by differential thermal analysis (D.T.A.) and X-ray diffraction. EXPERIMENTAL Sample Preparation. Starting materials, SnTe and SnSe, were prepared by the direct fusion of commercially available high-purity (99.999 pct) elements. Stoichiometric amounts of each couple Sn-Te or Sn-Se were weighed into a clear fused silica ampule. After evacuation to a pressure below 10-3 mm Hg, the am- pule was sealed, and annealed at 900°C for 5 hr. The melt was quenched in water. X-ray analysis confirmed the formation of a single phase of SnTe or SnSe. The other samples, SnTel-,Sex were synthesized from these SnTe and SnSe by mixing them in the required ratio, followed by annealing at 900°C and quenching. These samples were used directly for D.T.A. For X-ray measurements, samples were annealed at 700°, 600°, or 500°C for 100 hr or at 300°C for 150 hr, and then quenched in water. It was found that the lattice constants of the SnTe phase annealed for 150 hr at temperatures above 500°C did not differ from those annealed for 100 hr at the same temperatures. However the X-ray phase analysis showed that at 300°C the annealing for 150 hr was necessary to attain a true equilibrium state. D.T.A. The solid-liquid equilibrium temperature was determined from D.T.A. measurements. The sample was sealed in an evacuated silica tube and molybdenum powders sealed in an another tube were used as a reference material. The sample and the reference tube were placed in a nickel block and were heated from room temperature to 900°C at a rate of 3°C per min and then cooled down at the same rate to 600°C. Thermocouples for these measurements were Pt-Pt. Rh (10 pct) and the error of temperature measurements was within + l0C. D.T.A. curves were obtained on a two-pen recorder and an automatic controller (PID type) was used for the program of heating and cooling. When temperature reaches the solidus from the low-temperature side, there appears an endothermic peak. The solidus temperature was determined by extrapolation of the straight portion of the starting flank of this peak to the base line. In a similar way, the liquidus temperature was determined from an exothermic peak on D.T.A. cooling curve. In the case of supercooling, if any, its degree can be estimated from the magnitude of the abrupt temperature rise. X-Ray . X-ray powder patterns were taken by a diffractometer using CuK, radiation. Since the SnSe crystal is cleaved easily, the powders become flaky when SnSe-rich samples are ground in an agate

Jan 1, 1969

By S. N. Singh, M. C. Flemings

Results are presented of a detailed study on the combined influences of ingot dendrite am spacing and thermomechanical treatments on the structure and solution kinetics of high --purity cast and worked 7075 alloy. Solution kinetics were found to depend sensitively on ingot dendrite am spacing and on details of therrnomechanical processing, including amount of reduction and extent of&apos; solution treatment before reduction. An approximate analysis is given for rate of solution of nonequilibrium second phase in the cast and worked structres; results of the analysis are compared with experiment. MICROSEGREGATION in high strength aluminum alloys manifests itself as "coring" (composition differences within the primary aluminum-rich phase), and as interdendritic second phase. The mechanism of formation of the microsegregation is understood, and approximate prediction of the amount of second phase is possible for simple binary systems.1,2 When alloy elements or impurities are present in amounts less than their solid solubility at solution temperature, any phases forming from these elements are termed "nonequilibrium" and can be dissolved by appropriate solution treatment. The rate at which the nonequilibrium phases are removed depends sensitively on their spacing (dendrite arm spacing in the cast material, or band spacing in wrought material). When alloy elements or impurities are present in amounts in excess of their solubility at the solution temperature, second phase particles form an "equilibrium" second phase that does not dissolve in heat treatment and may, in fact, coarsen in such treatment. Usual commercial, high strength, wrought aluminum alloys contain nonequilibrium second phases that were not fully dissolved during ingot processing. They also contain equilibrium second phases resulting from impurities present in amounts greater than their solubility. As has been shown by Antes, Lipson, and Rosenthal,3 and will be demonstrated further in a subsequent paper by the authors,4 significant improvements in mechanical properties of high strength alloys can be achieved by reduction or elimination of these second phases. Methods of elimination are 1) to employ high purity materials to minimize amounts of equilibrium second phase, and 2) to employ suitable thermomechanical processing techniques to fully eliminate nonequilibrium second phases. Work reported herein comprises a study of selected thermomechani- cal processing treatments, and of their influence on solution kinetics of wrought high purity 7075 alloy. EXPERIMENTAL PROCEDURE Melting and Casting. The bulk of the work reported was performed on a single ingot of high purity 7075 alloy. The ingot was 4 in. by 4 in. by 8 in. high, uni-directionally solidified following a procedure previously described.5 The mold was heated to 1350°F before pouring the melt. The bottom chill was carbon coated stainless steel. Water was circulated through the chill after the melt was poured. The 7075 alloy was prepared from high purity virgin material (aluminum, zinc, magnesium) and from master alloys (Al-50 pct Cu, A1-15 pct Cr, A1-5 pct Ti). Final measured melt composition (wt pct) was: Zn Mg Cu Cr Ti Fe Si Al 5.70 2.28 1.35 0.18 0.15 <0.002 <0.012 bal Melting was done in a silicon carbide crucible; all tools were coated with zircon wash to minimize iron contamination; degassing was by bubbling chlorine through the melt. che-rmomechanical Treatments. Detailed studies were made on material taken from a location approximately 13 in. from the chill and 51/2 in. from the chill (i.e., from 1 in. thick slices taken between 1 and 2 in. from the chill and between 5 and 6 in. from the chill). Solution treatment was done at 860°F in an air-circulating furnace with a "bottom drop" arrangement to achieve minimum delay time between solution treatment and quench. Samples solution treated in this way were 2 in. by 2 in. by 1 in. Temperature of the quench water was approximately 10°C. Mechanical reduction was by cold rolling. Samples 11/2 and 51/2 in. from the chill were treated for 12 and 24 hr, respectively, before cold rolling. Reduction by cold rolling was then 4/1, 16/1, and 35/1. In each case, several intermediate anneals (1/2 hr at 860°F) were used to permit reaching the final thickness without cracking; two such anneals were used for the 4/1 reduction, five for 16/1, and six for 35/1. After working, materials were again solution treated for various lengths of time from 0 to 48 hr and quenched in water. Structural Measurements. Secondary dendrite arm spacings were measured using procedures previously described.&apos; For each measurement reported, five photomicrographs were first made at X75. Measurements were made of dendrite arm spacings in at least 20 different grains (grain structure was equiaxed). Grain size measurements were made by running a number of random traverses across photomicrographs of the samples and obtaining the mean lineal intercept. Measurement of the volume percent of second phase and porosity was done by quantitative metallography. A two-dimensional systematic point count was used

Jan 1, 1970

Jan 1, 1901

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

W. J. PARTON*—I have not had the opportunity to read this paper, and I do not have a written discussion. However, I thought it might be interesting for me to relate some of the experiences we had with equipment similar to the vibrating filter as described by the&apos; authors. At the Tamaqua flotation plant of the Lebigh Navigation Coal Co. approximately 40 tons per hour of froth concentrate carrying 60 pct by weight moisture are produced. The major problem encountered at this plant is the dewatering of this coal froth so that a satisfactory product can be sent to market. In the original design of the plant a centrifuge of solid bowl type was included for de-watering this material. The centrifuge did not work out as well as we had hoped. High maintenance costs and moisture content in the cake were obtained. A Robbins dewatering screen was installed at a later date with the idea of using it in conjunction with the centrifuge. The froth concentrate from the flotation cells was fed directly to the Robbins dewatering screen. The cake from the screen carried approximately 55 pct of the feed solids. Moisture in the cake was approximately 24 pct by weight. The underflow from the screen carried 45 pct of the feed tonnage at about 80 pct moisture by weight. The underflow product was then pumped into the centrifuge with the idea of using the cen-trifuge for recovering the tonnage lost through the screen. This circuit did not operate as satisfactorily as we expected. The only benefit derived was in the reduction in the power consumed by the centrifuge. The maintenance on the centrifuge was approximately the same as previously. The next step in our experiments was to pump the underflow from the screen into a cyclone thickener which was mounted directly over the vibrating screen. This thickener increased the concentration of the solids to approximately 60 pct by weight and dropped the mate- rial back 011 the filter cake which had formed toward the discharge end of the screen. Unfortunately, the screen was not capable of handling this additional tonnage, and our experiments stopped at that point. We have been considering installing a second screen to make possible the complete mechanical dewatering of this product by the use of the dewatering screen and the cyclone thickener. Another possibility under study is to pump the underflow from this screen to a thickener which is available in the flotation plant, and to combine this thickeued underflow with the original feed going to the screen. Again, however, a second dewatering screen will be required to handle the total tonnage. 0. R. LYONS*—I had an opportunity to read this paper ahead of the meeting, and I did a little pencil engineering on it. As Mr. Bertholf said, it is very difficult to make a comparison and to carry the results of work at one plant over to what might be expected at another. What I did was to find information on filtering operations more or less comparable to the type of operation that Mr. Bertholf has with his vibrating filter. The only information that I was able to find was for drum type filters, and I found the operating characteristics of the vibrating filter and the drum type filters were very similar. The moisture contents of the cakes were almost identical. The output per square foot was about the only way that I could compare their capacities— using square foot of screen area against square foot of filter area—and I found the capacity of the vibrating filter to be slightly greater per unit area than the capacity of the drum-type filters. W. H. NEWTON†—Do I understand that the only escape for the solids is by overflowing the thickener? That is, does the filter have a chance to recover all the solids except that lost in the thickener overflow ? W. M. BERTHOLF (authors&apos; reply)— Actually, the only escapc from that part of the circuit is over the top of the thickener. There are other places the fines could be lost in the washery. but once they get into that part of the circuit, they must go over the top to escape. W. H. NEWTON—I would like to ask Mr. Lyons if, in the study of rotary filters, he has any basis for comparison of operating costs? 0. R. LYONS—No, I had no information on costs. The only information I was able to find was on screen size, moisture content, and tonnage output per unit area. W. L. McMORRIS*—Are you wasting that overflow water or re-using it? W. M. BERTHOLF—Right now, we are not re-using it. D. R. MITCHELL† —What is the approximate per capita cost of one of these units? W. M. BERTHOLF—It appears to be somewhere in the neighborhood of $200, for the screen. W. H. NEWTON—The cost would be about $2500 for the complete unit including the vibrating power unit. G. A. VISSAC‡—I do not like to come on the floor after I have been talking so long, but I thought you might be interested in our experience in dewatering, as well as drying our very fine coals. We have used both centrifuge and vibrating screens. The type of vibrating screens we have used in Canada are called the Zimmer. That is a screen of German construction, and I guess it is along the same lines as the dewatering screens you are using now. We use wedge wires, and the minimum size opening is a quarter of a millimeter. In our experience, the cheapest way is still a dewatering bin. A dewatering bin takes 48 hr to do work that takes 20 min in a dewatering screen. We use old wedge wire from our driers which we cover with brattice cloth, and

Jan 1, 1950

By A. E. Lee

THE Blackwell Zinc Co., Inc., a subsidiary of the American Metal Co., Ltd., operates a horizontal retort zinc smelter at Blackwell, Okla. The plant has 14 furnace blocks of 800 retorts each, fired with natural gas on a 48 hr cycle. Over 13,000 tons of zinc-bearing material, chiefly sulphide flotation concentrates, are treated monthly to produce slab zinc and high lead-cadmium fume. In 1942 a program of rebuilding and modernizing the smelter was started. By 1947 furnace smelting capacity had been increased to a point where roasting and sintering facilities were inadequate, and it was necessary to purchase oxidized materials to supplement sinter production. The seven 210 ft Ropp roasters and three 42 in. x 44 ft Dwight-Lloyd machines then in use had been in service at least 20 years and were in need of major rebuilding. Thus it was entirely practical to consider all new equipment and a change of method rather than rebuilding and repairing obsolete units. A study of the problem indicated that roasting as such could be eliminated and roasting and sintering accomplished in one step by a modification of the Robson process,&apos; which had been used since the early 1930&apos;s by the National Smelting Co., Ltd., at their plants at Avonmouth, England, and Swansea Vale, South Wales. Francis P. Sinn, General Manager, Zinc Smelting Operations, The American Metal Co., Ltd., who was familiar with the practice in England, suggested the use of one large machine for the entire operation from concentrate to sinter. One step sintering appeared to best meet Blackwell&apos;s plant requirements and indicated substantial savings in labor, gas, coal, and repair costs. Choice of Machine Size The sinter machine size was set at 12x168 ft for a rated capacity of 540 tons per day. This tonnage, produced on a five day week, would meet the seven day requirements of the 14 furnace blocks. The one large machine was quoted at a lower cost than two or more 6 ft wide machines of similar total capacity. Further, the larger machine could be housed in a smaller structure and only one set of equipment for charge preparation and delivery and for disposal of sinter cake was needed. One machine on a five day week made possible a concentration of the skilled operating personnel and required less men than a plant including two or more machines and related equipment circuits. Fewer units of equipment meant less maintenance, and the two down days weekly allowed ample time to repair and, if necessary, to make up lost production. Experience had indicated better sintering quality and rates with larger masses of material, not only on wider machines, but also in deeper beds. The ratio of windbox perimeter to area for the 12x168 ft machine is 0.179, compared to 0.353 for a 6x102 ft machine and 0.617 for a 42 in. x 44 ft machine. This meant less air leakage with resulting fan power savings and less spoilage of charge along the pallet sides. Performance Initial operation of the new sinter plant was made in November 1951 and regular production attained late in December. The average product sinter output during 1952 and the first half of 1953 has been 18.2 tons per hr. The average for one month has been as high as 22.4 tons per hr. Considerable experimenting with varied operating conditions accounts in part for the below capacity — 24 tons per hr — average output, and work to further improve production rate continues. A typical sinter analyses is 66.0 pct Zn, 0.3 pct Pb, 0.1 pct Cd, 0.3 pct S, 8.0 pct Fe, 2.0 pct SiO,, 0.8 pct CaO, and 0.2 pct MgO. Use of this material has made possible increases in furnace burden and improved furnace operation over the former practice using sinter made from Ropp roasted concentrates. Better lead and cadmium elimination in sintering has permitted the furnace production of slab zinc lower in lead and cadmium. Anticipated economies of operation have largely been gained. The sinter plant is operated by seven men per 8 hr shift — one head operator, three equipment operators and three sweepers — plus one oiler on day shift only. While it has been necessary at times to operate seven days a week to produce the required sinter tonnage, the five day work week usually has been adequate. Consumption of natural gas for sinter bed ignition is 200,000 to 300,000 cu ft per day. Green Ore Sintering Practice The 30 to 31 pct sulphur content of the —200 mesh zinc concentrates is the fuel used to sinter the charge, no coal addition being required. In the feed to the machine, sufficient concentrates are added to crushed return sinter fines containing 0.3 to 0.5 pct sulphur to produce a charge averaging 5.0 to 6.5 pct sulphur. Since the return sinter used in Blackwell&apos;s practice is varied from — 1/2 to — 1/8 in., the actual sintering mixture of fine sinter and concentrates is somewhat higher in sulphur. The coarser sinter particles are too large to resinter and merely aid porosity in the sinter bed. The ratio of concentrates to return sinter in the charge ranges from about 1:4 to 1:5.5. Variations are based on the appearance of pried up bed sections, bed exit gas temperature trends, windbox suctions, and return sinter size. Sufficient sulphur must be used to obtain fritting of the charge into a soft sinter cake and to aid in the elimination of lead and cadmium. Excessive feed sulphur will result in partial slagging of the cake impairing porosity and prolonging sintering time.

Jan 1, 1954

By E. B. Mayo

David LeCount Evans (Consulting Petroleum and Mining Geologist, Wichita, Kans.)-—Not only E. B. Mayo but also W. C. Lacy, who apparently urged the preparation of this analysis, is to be commended. Regional thinking of this type is needed to assure future success in the never-ending search for new mineralized and petroliferous districts. As is usually the case, here is a regional study that will be read by the mining geologist alone. It is ironic that several of the trends established in this study have suggested themselves in northern mid-continent, detailed, and regional studies. These, where established, have offered new keys to petroleum exploration and have provided a possible basis for unraveling a number of broad generalities. The oil geologists, active in Colorado, Kansas, and Oklahoma, would find much food for thought in Mr. Mayo&apos;s projections. To be more specific: 1) The parallelism between E. B. Mayo&apos;s Texas Lineament and the Amarillo Uplift, the Wichita Complex and the Arbuckle Complex of the Texas Panhandle and Southern Oklahoma is viewed with interest and appears especially significant when compared with the similar northwest trend of the Central Kansas Uplift, a major trend of production. 2) Considering the various northeast zones of Fig. 2, and with particular reference to Mayo&apos;s C-C, the Jemez Zone is on direct line with one of several northeast-southwest controls which the present writer has been using with some success in Kansas subsurface correlations. Considering zones of shearing, with no apparent vertical displacement, but suggesting strike-slip movement, because of the staggered effect on other features which cross such trends, Mayo&apos;s philosophy presents regional possibilities for lines of weakness, considered to this time of only local significance. 3) And, finally, in an area as distant from the Southwest as central Kansas, the north-south trends of the Fiarport-Ruggles anticline, the Voshel-Hol-low Nikkel-Burrton structures, the Dayton to Stut-gart trend, the north, slightly east trend of the Ne-maha structural complex, and others all seem to approach the north-south alignments, a through f, of Mayo&apos;s Fig. 3. Mayo&apos;s employment of structural intersections to pinpoint crustal weakness, to localize igneous activity and its accompanying mineralization is not, perhaps, a new concept, but it is a 1958 model, produced by tools improved from the ever-increasing accumulation of geological observations. The use of intersecting trends in petroleum geology is not a new idea, since much production in earlier days was encountered via the straight line projections of established trends to centers of intersection. A tragedy in this age of specialization is that iron curtains have been raised between groups, all seeking raw materials, all acolytes at the altar of structural geology, but all smugly content in and protected by the ivory towers of petroleum geology, engineering geology, mining geology, and geophysics. Mayo presents basic ideas which can stimulate mid-continent structural thinking and, in the case of cen- tral Kansas. he provides a key to replace the broad and overworked simple monoclinal, sinkhole-dotted, Karst topography credo, which is not finding its share of new oil in a state where the declining discovery ratio is disconcerting. The American Association of Petroleum Geologists would do well to add E. B. Mayo to its list of Distinguished Lecturers. Evans B. Mayo (author&apos;s reply)—In reply to David LeCount Evans&apos; comments, it is pleasing to learn that some of the elements discussed in my paper may interest petroleum geologists as well as mining geologists. This should not be surprising, however, because the lineaments make up the framework of the continent, and the oil-bearing sediments must reflect to varying degrees adjustments of basement blocks along their boundaries. A further possibility that petroleum geologists must have considered is that the slow escape of heat from buried lineaments and their intersections has aided the separation of oil from the sediments and started the migration into traps. Regarding the specific points listed by Evans, the following are suggested: 1) The branch of Texas Lineament marked 1&apos; (Fig. 3) is thought to extend eastward through the Capitan Mts., New Mexico, through the long Tertiary dikes east of Roswell, and beyond via the Matador and Electra ranges of the Red River Uplift, Texas. Its further continuation might be the eastern flank of the Ouachita Fold Belt. The Amarillo-Wichita-Arbuckle zone of uplifts appears to continue east-southeastward the Spanish Peaks belt (3-5, Fig. 3). The northwest-trending Central Kansas Uplift would not belong to the above set, except insofar as the Central Kansas Uplift is traversed by west-northwest folds, possible continuations of the Uinta belt (5-5, Fig. 3). 2) The possible continuations into Kansas of the Jemez zone are new to me and are most welcome suggestions. 3) Most of the nearly north-south Kansan structures mentioned by Evans are unfamiliar to me, but the Nemaha Uplift itself appears to be part of a very pronounced structure traceable from the Cerralvo Fault Zone, south of the Rio Grande, through the Bend Arch, Texas, and the Nemaha Uplift, into the Pre-Cambrian of Minnesota (?). This nearly meridional zone is crossed and broken by the Rio Grande Embayment and by the Red River-Wichita Syntaxis. Petroleum geologists realize the economic importance of these features. Perhaps it is inevitable that some papers of general interest be buried in the journals of specialized groups. Moreover, papers dealing with regional, or lineament, tectonics and its applications to exploration for economic mineral deposits are as yet few in the American literature. The opportunity to advance this field is open to all those who are not ultra-conservative and who have a lively curiosity, plenty of patience, and not too many business restrictions. In conclusion, much appreciation is extended to D. L. Evans for his comments.

Jan 1, 1960

By C. Altstetter

THE work herein reported is restricted to the carbides which occur in quenched and tempered AISI 43XX steels with carbon contents up to 0.40 pct and silicon additions of up to 3 pct. In view of the instability and extremely small size of the carbides formed at low tempering temperatures, the technique for successfully preparing specimens for X-ray diffraction will be outlined. The alloys listed in Table I were obtained through the courtesy of the United States Steel Corp. in the form of 1/2-in. rounds forged from 100 lb. induction furnace heats (except for 4337 which was a commercial heat). The stock was normalized and then swaged and drawn to 15 mil wire with anneals at 1200F between passes. The wire was austenitized for 45 min in evacuated vycor capsules and quenched into iced brine with simultaneous smashing of the capsule. Tempering was done in air with a water quench after tempering. The carbides were extracted in a simple cell using a solution of 1M KC1 and 0.5 pct citric acid with an initial current density of 0.1 amp per sq cm. One end of a short length of wire was immersed in the solution, and the current at constant voltage was noted as a function of time. After about an hour the current dropped sharply because of the decrease in specimen cross-section. At this point it was found that the dissolution could be stopped and that the very fine wire which then resulted was just large enough to permit handling of the extracted precipitate still clinging to it, yet so small that it diffracted and absorbed only a negligible amount of the X-radiation. This rod of residue with a convenient handle of undissolved wire was rinsed in distilled water. alcohol, and acetone. Then it was dipped in a thin solution of cellulose-acetate cement and dried in vacuum. The resulting specimen was straight, uniform in density, easily handled, but most important, was completely sealed and never exposed to air. Furthermore, the residue had never been subjected to strong acids or rough handling such as in the extraction-replica technique or in the complete extraction to a powdered residue. It was found that improperly coated specimens were pyrophoric, turning to oxide with a dull red glow as they were exposed to air and yielding patterns of Fe2O3 and Fe3O4. The steels containing 3 pct Si were especially difficult to prepare for this reason. The specimens were put in a 57 mm Straumanis camera with double pinholes or slits and irradiated with filtered-chromium radiation. Readable patterns were obtained in less than an hour. A preliminary finding of some note was that for both tempered and as-quenched specimens of steels 4337 and 4337 (1.5 Si). M23C6 patterns were found along with the patterns of other constituents of the residues. This result was somewhat surprising in that previous investigators had reported that this carbide did not appear in a 0.38 pct C, 0.48 pct Mo steel1 or in chromium steels of less than about 10 pct Cr.2 Although the total amount of carbide-forming alloying elements is less than 2 pct, due to their mutual interaction and the action of the plastic deformation in promoting equilibrium, this carbide was able to form even in the steel containing 1.5 pct Si. M23C6 was not detected in the 4337 (3.0 Si) steel and the lower-carbon steels were not investigated in this condition. It is very likely then that the steels studied herein underwent a fourth stage of tempering during the anneals at 1200°F. This result has significance in that even a small amount of undissolved M23C6 in a low-carbon, low-alloy steel would exert a large effect on its hardenability. Its presence would also influence the mechanical properties by decreasing the carbon content of the matrix. Annealing in vacuum for 1 to 4 hr in the austenite field removed all traces of MZ3C+ The results on carbide precipitation during tempering, summarized in Table I, are in agreement with those of Klingler et al.3 for the higher carbon steels. For the AISI 4337 steels it is noteworthy that in the steels with added silicon the E carbide persists to longer times and higher temperatures and that silicon delays the formation of cementite. The results for the lowzr-carbon steels parallel those of the higher-carbon grade. The appearance of E carbide in the AISI 4315 is significant. There is considerable disagreem-nt in the literature as to whether this carbide forms in the tempering of steels containing less than about 0.2 pct C. Following the detection of E carbide in a 0.18 pct C plain-carbon steel,4 its occurrence in a steel containing chromium and molybdenum should be expected. The fact that the low-carbon steels have the same carbide-precipitation sequence as the high-carbon steels has bearing on the larger problem of the exact tempering reactions in all steels. Following the suggestion of Roberts et al.,&apos; the first stage has been generally assumed to result in a metastable equilibrium of c carbide and martensite of about 0.25 pct C. From this it was concluded that a steel having less than 0.25 pct C should then be under-saturated with respect to c carbide and should not precipitate this carbide upon tempering. In view of the experimental findings of c carbide in steels hav-

Jan 1, 1962

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951