Search Documents

By B. Ramaswami, U. F. Kocks, B. Chalmers

The latent hardening of silver and an Ag-Au alloy was investigated by lateral compression, overshoot in tension and cormpression, and the stability of multiple-slib orientations. The latent hardening of a secondary slip systenz depends on its relation to the primary slip system. For most secondary slip systems the latent hardening is larger for Ag-10 at. pct Au than for pure silver. The maximum increase in. flow stress on a secondary slip system over that of the primary slip system was 40 pct. The work hardening during the lateral-compression test on the latent system after prestress on the primary system is iuterbreted in terms of the preferential distribution of barriers to dislocation movement with respect to the active slip system in work-lzardened fcc crystals. The work-hardening in fcc crystals is mainly due to the dislocation interactions and the barriers to dislocation movement formed as a result of reactions between dislocations of different slip systems. The operation of sources on the latent system depends on the flow stress of those systems; hence, the increase in flow stress of a latent system due to glide on an active system, which is called latent hardening, is an important element in understanding the phenomenon of work hardening. The problem of latent hardening has attracted the attention of many investigators in the past. For example, a theoretical study of the elastic latent hardening of the latent systems due to glide on an operative system has been made by Haasen' and ~troh. These calculations, however, neglect the stress required for the intersection of forest dislocations by the glide dislocations, a factor which would be important for producing macroscopic strains on the secondary slip systems. The importance of this factor will become evident from the results presented here. Attempts have also been made to determine the latent hardening of different slip systems by experimental means by the methods summarized in Table I.3-9 The experimental methods used have been subject to certain limitations. For instance, in the method used by Hauser,9 frictional constraints between the specimen and the compression platen were not eliminated by proper lubrication (see Hos- ford10). Secondly, with the exception of Kocks,6 Hauser,9 and Rohm and Kochendorfer,11 latent-hardening studies have been made on only one of the slip systems, i.e., on either the conjugate or the coplanar slip system; hence, extensive results are not available on the latent hardening of different slip systems in the same materials, with the exception of aluminum.6 It was therefore decided to study the latent hardening of the conjugate, critical and half-related slip systems in silver. Similar experiments were done in Ag-10 at. pct Au to study the effect of solute (gold) on the latent hardening of silver. Lastly, indirect evidence can be obtained by a study of the orientation stability of crystals of multiple-slip orientations in tension and compression. This method has been used by Kocks6 to supplement his studies of latent hardening in aluminum. Similar studies were made at room temperature in single crystals of silver. EXPERIMENTAL PROCEDURE The single crystals of the desired orientations were grown and the tensile test specimens were prepared as described in Ref. 12. The compression tests were made on 1/4-in.-cube specimens. The specimens were cut from single crystals, in the Servomet spark-erosion machine.13 The two cut surfaces were planed using the lowest available planing rate in the machine to minimize the deformation layer. A brass strip was used as the planing tool. This method of preparation ensured plane parallel faces for the compression tests. The deformed material was removed by prolonged etching in a weak etching solution. A weak etching solution was used to prevent pitting of the surfaces and to ensure uniform etching. About 25 to 50 µ of material were removed from all faces by the etching treatment. The specimens were then annealed for 24 hr at 940°C in oxygen-free helium and cooled in the furnace to room temperature over a period of 7 hr. After annealing, the orientation of the specimens was determined by Laue back-reflection technique to make sure that no recrystallization had occurred on annealing. The compression-test technique and setup are described in Ref. 14. The Laue back-reflection technique was used to study the overshoot in tension, the overshoot in compression, and the stability of the axial orientation in tension and compression. The tests were interrupted after every few percent strain to determine the axial orientation. In investigating the overshoot in compression, the operative system was determined by studying the asterism of the Laue spots.

Jan 1, 1965

By J. B. Newkirk, A. H. Geisler

The alloy Cu4Pd has a disordered face-centered-cubic structure when quenched from temperatures between 478ºC and the melting point (about 1100°C). Below 478ºC an ordered phase is stable. The results of a Debye-Scherrer X-ray analysis indicate that the ordered phase has a tetragonal unit cell described by the space group C24h — P42/mt with 2 Cu in 2a, 2 Cu in 2f, 4 Cu in 4j (x = 0.2, y = 0.6), 4Pd in 4j (x = 0.4, y = 0.2), and 8 Cu in 8k (x = 0.1, y = 0.3). The orientation relationship between the face-centered-cubic phase and the ordered tetragonal phase is given by: [100],,. // [130]al,. COO1Ia.d.//COO1I,,.. • The behavior of Cu,Pd is typical of ordering alloys except that the transformation is very sluggish. The increase in hardness and the microstructural and X-ray diffraction effects are interpreted in terms of coherency strains caused by the ordering. AN anomalous construction in the Cu-Pd phase diagram (Fig. 1) was reported in 1939 and has been allowed to stand without further published attention since that time. The odd figuration about the composition 10 to 27 atomic pct Pd is derived mostly from the work of Jones and Sykes.1 Evidently several features of this binary system require further study if the constitutional forms are to be well understood. The present paper includes a study of one of these features, that is, the crystal structure of a single ordered alloy containing nominally 20 atomic pct Pd. This choice of composition was suggested by the work of Harker and associates who determined the structure of Ni4Mo2 and Ni4W.3 The nature of the ordering process in Cu4Pd was studied also by observing the hardness, microstructure, and Debye-Scherrer patterns of specimens which had been aged at various temperatures after quenching from an initial disordering treatment. Experimental Methods A 20 gram ingot of Cu4Pd was made by melting spectrographically standardized copper from Johnson, Matthey, and Co., and commercially pure (99.5 + ) palladium in an argon-filled quartz tube. Chemical analysis showed that the ingot contained 80.0 atomic pct Cu. The ingot was rolled about 60 pct to a strip 0.060 in. thick and was homogenized for 16 hr at 950°C in low pressure argon. Rods cut from the rolled strip were worked into wire 0.015 in. in diameter, and specimens for hardness and microscopic examination were cut from the remaining strip. All specimens, with the exception of some of the wire, were given an initial disordering treatment by heating for 16 hr at 950°C, followed by water quenching. A 10 cm length of as-drawn wire was water quenched after being held in a temperature-gradient furnace4 for 89 days. Room-temperature Debye-Scherrer photograms were then made at points along the wire to determine the temperature below which the ordered phase was stable. Although the accuracy of temperature determination in the gradient was only about ±10 °C, the temperature gradient was sufficiently gradual that the sensitivity was much better and locations which had differed by as little as 1°C could be distinguished. An analysis of the crystal structure of the well ordered alloy was made by X-ray diffraction using a specimen cut from this wire. The change of Debye-Scherrer pattern as ordering progressed was studied by using isothermally aged samples of initially disordered wires. The wires were sealed under low-pressure argon in small quartz tubes for heat treatment. After the aging treatment, the tubes were quenched in water and photograms were made at room temperature in a 10 cm diam camera using filtered Cu kX. (A = 1.540511) Hardness was measured on a Vickers hardness tester using a 10 kg load and 2/3 in. objective lens. Reported values are the average of at least three impressions made on flat specimens 0.060 in. thick. After the hardness of a heat-treated sample had been measured, it was resealed in low-pressure argon and returned to the furnace for continued aging at the same temperature. In this way, two samples served for all aging times at each temperature. Hardness specimens which had been aged 500 hr or more were used for metallographic examination after the final aging treatment. A dilute potassium-dichromate etching solution was used.

Jan 1, 1955

By S. S. Arentz

PLANNED health and safety programs have become an essential part of American industry because such programs lead to increased operating efficiency, improved labor relations, better public relations, and to substantial savings in compensation insurance. Those of you who have had the unpleasant duty of informing the wife or widow of one of your men of his serious injury or death while on the job, know that all the benefits of a successful safety program do not show on the balance S. S. ARENTZ, Member AIME, is General Superintendent, Nevada Operations, Combined Metals Reduction Co., Pioche, Nevada. AIME San Francisco Meeting, February 1949. TP 2741 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Jan. 6, 1949. sheet. These programs are of particular importance to the mining ,industry because mining's reputation as an unusually hazardous industry and the commonly isolated location of mining operations tend to focus attention on these problems. Description of Operations: Before proceeding with a discussion of our health and safety programs at Pioche, it may be proper to give a brief description of Pioche and of our operations there. Pioche is one of the early Nevada mining camps. It was founded shortly after the discovery of high grade silver ore in 1863 and mining has continued with more or less regularity to the present day. In an era of lawlessness, Pioche was notorious. The story persists that 75 men died with their boots on before one died a natural death, and old payroll records show that nearly as many gunmen were employed to stand off claim jumpers as there were miners working the mine. That was probably as close to a safety program as the times permitted. Pioche is situated in southeastern Nevada on the main highway between Ely and Las Vegas. The camp is on the flank of "Treasure Hill," near the original silver discovery, at an elevation of about 6000 ft. The present day population of about 2000 is primarily dependent upon the mines of the area, although Pioche also serves as the county seat of Lincoln Couqty and as the center of the surrounding livestock industry. The camp is served by a branch of the Union Pacific Railroad and receives power from the generators at Hoover Dam. The Pioche operations of the Combined Metals Reduction Co. were started in 1923 when the first complex lead-zinc ore was shipped to the company's mill at Bauer, Utah. The modern mill at Pioche was completed in 1941. The operations are medium sized in the nonferrous field, employing an average of 350 men in the mine, mill, and related works. The complex lead-zinc ore is mined from replacement deposits in a comparatively flat, extensively faulted, limestone horizon. Mining methods vary from stull-supported open stopes to filled square-set stopes. The thin bedded limestone and shale overlying the ore is allowed to cave as areas are mined out and caving frequently follows closely upon ore extraction. The relatively heavy ground and the numerous faults add to the problems of safe mining. The mine is well mechanized and the mill and surface plant are modern and well equipped. Labor is organized in a C.I.O. union and labor-management relations have been unusually harmonious. During most of the period since 1923 a competent supervisory staff worked to reduce safety hazards but the primary responsibility for safety rested on the individual workman. Accidents happened and all too frequently they were regarded by all concerned as unavoidable. In October 1939, the late Robert L. Dean became superintendent at Pioche. Most of his previous experience had been in the fields of iron and coal mining and from that experience he brought the concept that no accident is unavoidable. Many of the features of our present health and safety programs were initiated by Mr. Dean during his term as superintendent. Health Program: Our health program centers in Dr. Q. E. Fortier and his new, well-equipped, and well-staffed, modern hospital in Pioche. The program starts with a thorough pre-employment physical examination and is followed by yearly re-examinations at the expense of the company. The Pioche Mutual Benefit Association, to which all Pioche mine operators and employees belong, pays benefits covering hospitalization and surgery expense incurred by employee members and their families. The Association is governed by a board of directors elected by its members. The mine operators of the district donated the original capital and pay the monthly dues of the employee members. The employees pay the dues covering members of their families. Though not strictly a part of the

Jan 1, 1951

By C. H. Lee, R. Maddin

The diffusion rates of Ag110 in cylindrical single crystals of 99.99 pct pure silver, which were twisted to a total surface strain of 0.91 during the diffusion anneal at temperatures of 700°, 750°, and 800°C, were measured by sectioning and counting as a function of the strain rates. The ratio of the diffusivities while strain ing occurred (Ds) to those of the unstrained specimens varies linearly with the rate of twisting and attains values of near 100 for some of the rates used. The enhancement of diffusion by straining is inversely proportional to the temperature at which diffusion occurs. The activation energy for diffusion determined from the data plotted as a function of the rate of twisting decreases from 44,090 cal per mole for zero strain, to an asymptotic value of 21,000 cal per mole for the strain rates of 22 x it)"5 see-l. Qualitative explanation for the observations based on the generation of nonequilibrium number of defects by plastic deformation will be discussed. The initial research of Buffington and cohen1 on self-diffusion in a iron under compressive stresses contributed the first quantitative information concerning the effect of applied stresses on diffusion. These results were obtained from radioactive Fe55 plated on one of the horizontal surfaces and compressing the large-grained specimens while the diffusion-anneal was occurring. Enhancement of diffusion of a maximum of seventeen times was observed provided the strain was in the tertiary stage of creep and the rates of straining were at least as fast as rates used in the normal creep test. On the other hand, Darby, Tomizuka, and Balluffi,2 extending silver single crystals up to 30 pct elongation at rates of straining about equal to those used by Buffington and Cohen observed no effect. For the past two and a half years at the University of Pennsylvania, we have been investigating the effect or torsional strains on self-diffusion in silver single crystals. If enhancement is observed as has been reported1 a possible explanation may be based upon the generation of fairly large numbers of point defects and their role in the diffusion process. Hence, since torsion should produce an abundance3 of point defects, an increase in the diffusivity while the crystal is being twisted should easily be measured. EXPERIMENTAL PROCEDURE Our research uses single crystals of silver prepared from 99.99 Ag by the Bridgman method in a dynamic vacuum of less than 25 µ. As grown, they are about 25 cm long by 1.3 cm in diam. The crystals are cut into 7.5 cm lengths with a jeweler's saw and machined to the shape shown in Fig. 1. The test length of the crystal is machined to a 0.6 cm diam after which the ends are masked and the test section etched heavily in HNO3, and KCN:(NH4),S2Os , solution to remove the effects of machining. The specimen is then annealed in vacuum for 8 to 12 hr at 800°C to remove any residual stresses and also to recrystallize the end sections. The recrystallized ends are necessary since experiments showed that single-crystalline end sections are too soft to transmit the desired torque. Occasionally, recrystalliza-

Jan 1, 1960

By Y. C. Liu, W. R. Hibbard

Based on pole figure data and microstructural observations, the re-crystallization orientation found in a copper strip previously cold-rolled 99.5 pct from a single crystal with an initial (110) [112] orientation may be described as: a 30° rotation, clockwise and counterclockwise, about octahedral poles of the cold-rolled texture such that all four poles function at a low annealing temperature (400°C) and only three of them function at high temperatures (500' to 1050°C). The relative intensity of deformation stresses on various slip planes can be correlated with the choice of poles affecting the rotations found in the recrystallized orientations. ROTATIONAL reorientation about the poles of densely packed crystallographic planes is an important characteristic of secondary recrystallization" in face-centered cubic metals. Some evidence has also been interpreted to show that such a rotational relationship also exists in the primary recrys-tallization process." In an investigation of the crystallographic relationships in the recrystallization process, several aspects of the experimental procedure should be considered. With cold-rolled poly-crystalline metal as the initial material, pole figure analysis leaves many ambiguities as to whether or not simple indices can be assigned to adequately represent its orientation. The rolling texture does not consist of a single texture. Minor deformation textures are usually present. In all cases, twin textures are present. The presence of both minor and twin deformation textures influences the recrystallization texture. In order to study the recrystallization mechanism, a more precise knowledge of the history of the material before recrystallization is necessary. Tensile deformation of a single crystal usually produces residual stresses which are concentrated at slip bands. Since the exact nature and orientation of these potential nuclei sites cannot be experimentally evaluated, the tensile deformation of a single crystal was not considered adequate. From the work done by Barrett and Steadman6 n copper, it appeared that cold rolling of a single-crystal specimen with an initial (110) [112] orientation would yield a specimen which fulfills the conditions required for the present investigation, namely a highly preferred single rolling texture with relatively homogeneous stress distribution. By investigating the recrystallization texture of this material, additional details of reorientation during recrystallization might be obtained. The purpose of this paper is to describe such an investigation. Experimental Procedure Copper used in the present investigation was cathode sheet with a purity of 99.94 to 99.97 pct. A copper single-crystal specimen, 1.15x0.8720x0.5322 in. (thickness) was cut from a cylindrical crystal which was grown in a Bridgman furnace.' The cutting was accomplished on a horizontal milling machine operated at a very slow speed to obtain a (110) plane in the rolling plane and a [112] direction in the rolling direction. The disturbed surface was removed by electrolytic polishing in dilute orthophos-phoric acid (H3PO4) with a specific gravity of 1.14, and a current density of about 1 to 2 amp per sq cm. A 3-min polishing was needed to eliminate the disturbed surface causing Debye-Scherrer rings and a 45-min polish produced a surface which yielded sharp rounded Laue spots. It was estimated that about 1/16 in. of metal was removed from the original cut surface. The final orientation of this specimen before rolling was about 2" from the (110) plane in the rolling plane with a [112] direction aligned in the rolling direction. It is possible that a small amount of strain was still present in the cube despite the fact that the Laue spots seemed very sharp. It is doubtful, however, that this would be an important factor after a subsequent rolling reduction of 99.5 pct. This copper specimen was rolled on a laboratory rolling mill with two highly polished rolls 37/8 in. in diam in the following manner: 0.010 in. per pass to 50 pct reduction, 0.005 in. to 70 pct, 0.002 in. to 85 pct and, finally, 0.001 in. per pass to a 99.5 pct reduction. Specimens, about 1 in. sq with the rolling edges intact, were cut from the rolled strip by the electrolytic method previously described, after the strip

Jan 1, 1954

By A. H. Daane, J. J. Hanak, F. H. Spedding

IN the study of the rare earths at the authors' laboratory, methods have been devised for preparing pure rare earths and many of their properties'- have been determined. Very little information is found in the literature on the properties of europium metal, which is the scarcest of the naturally occurring rare earths. Since we have recently prepared europium in appreciable quantities, we are in a position to describe the method of preparation and some of the properties of this metal. Metallic europium was first prepared by Klemm and Bommera who heated a mixture of potassium and EuC1, in vacuo, and obtained a mixture of europium metal and KC1. Trombe4 electrolyzed a molten salt bath containing NaC1, KC1 and EuC1, and collected the europium in a molten cadmium electrode as a cadmium alloy. The cadmium was subsequently removed by distillation but the properties of the residue indicate that pure europium was not obtained. McCoy attempted to prepare europium by distilling mercury from europium amalgam, but the stability of the Eu-Hg intermetallic compounds prevented a complete separation of the two elements. The preparation of very pure europium was suggested by previous work in this laboratory1,4 in which ytterbium and samarium were prepared by heating the respective sesquioxides with lanthanum metal turnings to 1450°C in vacuo and distilling ytterbium and samarium from the mixtures. Subsequently, small test experiments showed that europium also could be prepared by this method. A more detailed version of this method was used in the work described in this paper. Klemm and Bommer3 found by X-ray analysis that the metal had a body-centered cubic (bcc) structure with a lattice constant a, = 4.582 ± 0.00A (converted from kX units), with a bulk density of 5.30 g per cu cm. They found europium to be weakly ferromagnetic. Spedding, et al,7 found a,, to be 4.606 ± 0.001A. Barrett' showed that the bcc structure persists throughout the temperature range 5 to 3000K with a, at 5, 78, and 300°K being 4.551 ± 0.003, 4.551 ± 0.004 and 4.577 ± 0.001A, respectively. Trombe1 described europium as being iron gray, malleable, about as hard as neodymium and not attacked by cold water. He estimated the melting point to be between 1100 and 1200°C. Previous work in this laboratory indicated that the pure metal corrodes rapidly in air and reacts with cold

Jan 1, 1959

By H. H. Hausner, O. V. Roman, F. V. Lenel, G. S. Ansell

The radial shrinkage during sintering of cylindrical compacts and loose aggregates of copper powder was measured. It was found to be nonuni-form from top to bottom of the samples and to depend upon the method of supporting them. The non-uniformity is due to the effect of gravity forces during sintering. Since gravity has an effect in sintering without externally applied stresses, no sharp dividing line can be drawn between conventional sintering and hot pressing. RECENT investigations of the sintering behavior of compacts1 and of loose powder aggregates2 have indicated that forces, other than those arising from surface tension effects, may play a role in shrinkage. In compacts it was shown that residual stresses from the pressing operation influence shrinkage behavior. In loose powder aggregates gravity forces due to the weight of the powder affect the ratio of shrinkage in the vertical and the horizontal direction. The main effort in the work reported here was to show that gravity plays a role also in the sintering of compacts. A few additional experiments were made confirming the effect of gravity in the sintering of loose powder aggregates. EXPERIMENTAL PROCEDURE Compacts and loose powder aggregates were prepared from irregularly shaped, electrolytic copper powder. Prior to use the powders were reduced 30 min at 400°C in dry hydrogen to remove surface oxides. Then the -325 mesh size fraction was separated from the -100 + 325 mesh fraction. The compacts were pressed at a pressure of 10,000 psi from 50 g of powder in a hardened steel die, 1 in. in diam. The height of the compacts was 0.725 i 0.005 in. An effort was made to get as uniform a green density distribution in the compacts as possible. The walls of the die were lubricated with a suspension of 3 pct of zinc stearate in acetone and the compacts were pressed using double action by first pressing the powder at 1200 psi with the die barrel supported, then removing the sup- ports and pressing to final pressure of 10,000 psi with the die barrel floating. The pressure was maintained for 10 sec. The compacts were sintered at a temperature of 925°C, generally for 1 hr. In order to maintain uniform temperature they were sintered in boats made from cylindrical copper blocks. The blocks were 2 in. in diam, either 2 or 2 1/2 in. long and, split to form the body of the boat and a lid. The body of the boat contained a cavity 1 3/8 in. wide, 1 in. deep and either 2 1/8 or 1 3/8 in. long. The longer cavity accomodated two samples, the shorter one only one sample. The uniformity of temperature distribution within the boats was checked with thermocouples welded to the top and bottom of the samples. The maximum variation between top and bottom temperatures was i 1/2°C. The actual sintering temperature was held constant within ±2°C. In order to determine the effect of gravity forces, i.e., the weight of the compacts, upon shrinkage, they were supported in the following ways during sintering: a) Full Bottom Support. The compacts rested either on a flat alundum disk or on alundum powders. b) Partial Bottom Support. The compacts rested on a graphite cylinder, 0.3 in. in diam which formed a projection on a larger graphite disk. It is difficult to balance the compact on the small projection. To avoid having the compact tip, a small hole was drilled through the compact and through the graphite disk and its projection. The graphite disk was then suspended from the lid of the boat by a thin iron wire which passed through the holes in the disk and the compact. c) Top Support, first type. A hole 3/32 in. in diam was drilled diametrically through the green compact 3/16 in. from the top of the compact. The compact was sintered suspended from an alundum rod (thermocouple protection sleeve) inserted into the hole. d) Top Support, second type. A hole was drilled axially through the center of the compact. The upper part of the hole from the top surface of the compact one fourth of the way down was 1/16 in. in diam; the lower part of the hole from the bottom surface three fourth of the way up was 3/32 in. in diam. The compact was suspended from the lid of the boat by an iron wire passing through the upper part of the hole and then tied into a knot. The loose powder aggregates were made by filling l in. deep, l in. diam cylindrical graphite molds with -325 mesh powder. To achieve uniform density in all the loose powder aggregates, the powders were settled in the molds by placing the

Jan 1, 1963

By William H. Breeding

The remarks that follow are based substantially on experience covering 45 years, 80% of which has been in placer work, rather than on a review of available literature. Most commercial placers have been deposited by the action of water. The richer and more- difficult-to-mine placers are those in the headwater areas where gradients are steepest. The most lucrative placers are generally in inter- mediate areas where volumes are greater, fewer boulders are present, and gradients are from 3% to 1-1/2%. The higher volume, lower grade placers are in the lower reaches of river systems where gradients are lower. Where gold-bearing rivers have discharged into the sea, wave action can concentrate values on beaches, past and present. Most of the rich, readily accessible placers were mined by our forefathers. Current opportunities exist: (1) in remote areas where infrastructure has been absent in the past, or development has been prohibited by adverse ownership - political or commercial; (2) in deposits that could not be mined by equipment available to our forefathers; (3) in deposits unidentified by our forefathers; (4) where the-price-of-product/cost ratio is substantially better than in earlier years; or (5) a combination of those factors. When I entered the placer business in the late 1930s, and subsequently, a prevailing opinion believed that glacial deposits should be avoided as irregular in mineral content and composition, and unrewarding to explore and develop; yet an operator has been mining a fluvio-glacial deposit profitably for the past 17 years. Rich buried placer channels, of ten called paleo-channels were worked in the last century, generally by hand methods, and under conditions that would be unacceptable today. Exploration and mining equipment now available make some of these channels attractive targets. Well-known examples are in California and Australia. The formation of a commercial placer requires a source of valuable minerals. Above primary deposits, there may be eluvial deposits formed by the erosion of gangue minerals and the concentration "in situ" of valuable minerals. Down slope from these deposits are the hillside or colluvial deposits, and below them are the alluvial deposits of redeposited material. Most of the great placer fields of the world are the result of several generations of erosion and deposition. Well-known examples are in California and Colombia. Gold is a very resistant and malleable material, and gold placers may extend for 64 or 80 km (40 or 50 miles) along a river system. Platinum is less malleable, but is very resistant to disintegration. Diamonds are extremely hard, and (especially gem diamonds) may be found over great lengths of a river system. Cassiterite is less resistant to disintegration, and tin placers seldom extend over two miles without resupply from an additional source or sources of mineralizaton. Tungsten minerals are generally more friable, and within a few hundred yards of the source disintegrate to the point that they are uneconomical to recover. Rutile, ilmenite and zircon placers generally result from the weathering of massive deposits, and may be encountered over extensive areas; most are fine grained and durable. What does a geologist or mining engineer look for in placer exploration? The old adage to look for a mine near an existing mine is still valid. You need a source of valuable mineral. Then you require conditions for concentration, which means a satisfactory gradient and/or other conditions that will permit heavy minerals to settle. Nicely riffled gravel, often called a shingling of the bars, is conducive to placer formation. Coarser gravel is logically associated with coarser gold. Excessive clay and/or high stream velocities in narrow channels can carry gold far downstream and distribute it uncommercially over a large area. When material is extremely fine, in situ weathering and concentration become more important. Placers frequently occur distant from lode mines, and one must remember that in a larger watershed the exceptional floods that occur once in a hundred or a thousand years can move great quantities of material long distances. The carrying power of water is said to vary with the fifth or sixth power of its velocity. I am not ready to disagree with Waldemar Lindgren and accept that many commercial placers are substantially enriched by the chemical deposition of gold from solutions; however, I have seen crystalline gold in clayey material quite distant from known sources of primary gold that is dif-

Jan 1, 1985

By T. B. Massalski, Horace Pops

The occurrence and the temperature dependence of the athermal martensitic transformation in bcc Cu-Zn ß-phase alloys have been studied by cold-state microscopy, differential thermal analysis, and electrical -resistivity measurements. CinenzatograpJzic studies have shown that in the majority of alloys the transformation occurs in two stages. A thermoelastic phase, having a 'needlelike" uppearance, forms initially at the MS temperature and grows relatively slowly, chiefly in the edge (lengthwise) direction. With further cooling, an additional martensitic phase forms in rapid bursts at a lower temperature, MB, Both martensitic phases appear to have the same cry stallographic habit hut their modes of formation and growth are different. For the composition range studied, between 38.55 and 41.49 wt pct Zn, the habit planes determined are close to the (2.17 ,12) plane of the matrix, in general agreement with the prediction of the phenomenological theory of martensite formation. Thermal cycling through the MB temperature produces permanent plastic deformation in the 0 matrix and gradutally eliminates the typical burst phenomena. Thermal cycling only through the MS temperature produces no noticeable deformation effects. THE bcc Cu-Zn ß-phase alloys tend to become unstable at low temperatures and undergo a martensitic transformation,'-3 but the details of the transformation have not been completely determined. Following earlier metal log raphic and X-ray work1 Tichener and ever' determined by means of electrical-resistivity measurements the temperature range of the martensite transformation during continuous cooling. The transformation temperature was found to decrease with increasing zinc content, becoming about -130°C for an alloy containing 40.04 wt pct Zn. Hull and Garwood3 made a metallo-graphic study of Cu-Zn 0 brasses at temperatures down to -160°C, and determined the habit plane of the transformation product for an alloy containing 38.88 wt pct Zn. No metallographic studies below liquid-nitrogen temperatures have been reported. The martensitic transformation in ß-brass alloys can also be induced by deformation.1,4 The MD temperature corresponding to the appearance of the strain-induced martensite is some 300°C higher than that observed without strain, and hence the transformation can be strain-induced in alloys containing a higher percentage of zinc. Massalski and Barrett4 found that alloys containing as much as 51.89 wt pct Zn could be transformed into a faulted hcp structure by cold working at liquid-helium temperature. Because the details of the MS temperature and morphology reported in the literature are somewhat incomplete, a further study has been made of some of the above features. A design of a cryogenic stage for optical metallography5 has permitted the examination of specimens at temperatures down to 4.2oK, and the use of a cinecamera has permitted a study of the morphology on cooling and heating. The particular aim of the present work was to examine the transformation temperatures, the habit planes, and the growth morphology over a wide composition range in the ß-phase field. The effect of thermal cycling upon the subsequent transformation was also examined.

Jan 1, 1964

By ED. E. Furman, R. H. Atkinson

HITHERTO the practical creep testing of precious metals has received little or no attention. The only previous creep tests of precious metals have been made with wires under conditions such as to yield much more rapid rates of creep than in engineering tests.', ' Up to the present time the value of creep bars of adequate size, in the absence of real need for engineering data, has deterred investigators. However, the increasing use of platinum at high temperatures has demonstrated the need for reliable creep data for the guidance of engineers, especially those engaged in designing certain specialized chemical plant equipment. In order to supply this need, creep tests were conducted at 1382°F (750°C) on 0.290 in. diam specimens of platinum, 90 pct Pt, 10 pct Rh and palladium. The platinum was high purity, nominally 99.95 pct Pt. The 90 pct Pt, 10 pct Rh was of the same high quality as is used for making gauzes for the catalytic oxidation of ammonia. The palladium was also of high purity; two batches of palladium bars were tested, one deoxidized with calcium boride and the other with aluminum. Spectrographic examination of the palladium confirmed its good quality; the only significant impurities apart from the residual deoxidizers were traces of silicon and lead. Procedure The creep bars, which were furnished by Baker and Co. to our specification, were 6 ¾ in. in overall length with a 4½ in. (4 in. gage length) reduced section 0.290 in. in diam and had the ends threaded (?-NC16). It may be of interest that the bars were valued at up to $600 each. The specimens were supplied in a 50 pct cold-worked condition to facilitate attachment of the creep extensometer, which was of the push rod type. Because of the softness of the platinum and palladium, the extensometer rings were secured to the test section by means of circular knife edges instead of the usual pointed set screws. The extensometer rods extended through the bottom of the furnace and readings were taken with a 0.0001 in. "Last Word" dial gage fastened to the rods for the duration of the test. The bars were directly loaded by hanging weights from the lower specimen grip. All tests were conducted at 1382°F ± 2°F, and an effort was made to maintain the temperature gradient over the test section within 2°F. The ends of the furnace tube were packed with asbestos wool, which allowed a very slow circulation of air through the tube. Annealing was accomplished in the creep furnace before the load was applied. The platinum and palladium specimens were annealed at the test tem- perature for about 17 and 24 hr respectively; in the case of the rhodioplatinum it was found expedient to anneal for 1 hr at 1922°F (1050°C). Pilot samples cut from the same stock as the bars were used to check annealing procedures. Pertinent measurements of grain size and hardness were recorded. Results and Discussion The creep data obtained are given in Table I and the creep curves are plotted in Figs. 1, 2, and 3. Two platinum specimens, tested under a stress of 250 psi, had almost identical creep rates at 2000 hr, namely 0.000008 and 0.000009 pct per hr. A third platinum specimen, stressed at 400 psi, had a creep rate at 2000 hr of 0.000026 pct per hr; the reason for a rather sharp decrease in creep rate during the period from 1200 to 1600 hr is unknown. As it was thought that 90 pct Pt, 10 pct Rh would have a lower creep rate than platinum, the first sample was tested at 400 psi; however, the creep rate was approximately 50 pct greater. Microex-amination revealed that differences in grain size might be responsible for the unexpected result, as annealing at 1382°F developed an average grain diameter of 0.0021 in. in the rhodioplatinum specimen compared with 0.004 in. in the platinum bar. Annealing the alloy for 1 hr at 1922°F (1050°C) increased the average grain diameter to 0.0032 in. and materially improved the creep resistance, making it much better than platinum. A second specimen annealed at 1922°F (1050°C) and tested under a stress of 550 psi had a creep rate of 0.000022 pct per hr at 2000 hr, which was still substantially lower than that shown by the specimen annealed at 1382°F (750°C) and stressed at only 400 psi. In contrast to the creep behavior of the platinum and rhodioplatinum specimens, the palladium bars, whether deoxidized with calcium boride or aluminum, were characterized by high first stages of creep. However, after about 1200 hr of test, the creep

Jan 1, 1952

By C. K. Dumbauld, B. E. Morgan

The need for a low-strength cementing composition for use in well cementing is reviewed and results are presented of laboratory and experimental field tests of a modified cement having a controlled ultimate tensile strength of approximately 200 psi. The modified cementing composition may he prepared from either high early strength or normal portland cements by the addition of bentonite clay and a suitable agent for dispersing and controlling the set of the slurry. Substitution of the modified cement for conventional slow-set cements may give better completion results in many wells because the modified cementing composition has lower set strength, lower slurry density, and greater slurry stability than conventional cement slurrieh. The lower ultimate strength allows greater penetration with less shattering of the set cement when perforating casing and cement. The lower elurry density allows the placement of longer columns of cement slurry, and the greater slurry stability reduces the possibility of having an uncemented section caused by the settling of cement particles before the cement set.. INTRODUCTION High strength lias always been one of the accepted criteria of a good cement. During the early use of portland cement in well cementing. emphasis was placed upon securing cements with higher strengths. In 1931, Barkis' reported that, "Normal oil well cements have been improved to develop greater strengths and uniformity of product, which has aided in producing successful jobs in cementing the deeper strings." As long as most wells were completed by the open-hole method. the use of cements having high strengths seemed desirable, and there was no objection raised against high-strength cementing compositions. For a number of years. how-ever, the industry has been completing a large numher of wells by setting and cementing casing through productive horizons and then obtaining production by gun-perforating the section of casing opposite the desired interval. Although this method has been generally successful, difficulties have been experienced in some cases in completing or recompleting wells because of apparent lack of adequate penetration by the bullets through the casing and surrounding sheath of cement and into producing formation. In addition to the penetration trouble: there have been indications that fracturing and shattering of the set cement by perforating might he a contributing factor in causing the failure of some jobs to exclude water or gas from oil producing zones. The possibility that cements having high set strengths were contributing to til difficully ill obtaining satihfartory perforating results has re. ceived attention during recent years. Gun perforating tests conducted in 1944 showed that the depth of bullet penetration into set cement varied with the hardness of the cement, the greater the strength of the cement the less being the penetration. In 1946, Farris2 pointed out that high strength cements were not needed in well cementing, and in 1947, data published by Oliphant and Farris3 showed that set cement was perforated without shattering at approximately 150 psi tensile strength. whereas at approximately 300 psi tensile strength severe cracking and shattering occurred. Oliphant and Farris suggested that wells be perforated at the proper time interval after placement of the cement so that the set cement would have the desired strength. Several difficulties may he encountered in trying to perforate a cement job at the correct time to catch the tensile strength near 150 psi. The rate of strength development of different cements varies considerably. This fact is illustrated by results of tensile strength measurements presented in Fig. 1. These data show that at 175°F the tensile strengths of three conventional slow-set cements varied from 75 to 235 psi at the end of 12 hours. After 24 hours, the tensile strengths varied from 200 to 455 psi. The rate of strength development is affected. also. by the temperature of the forniation, and this adds to the uncertainty of perforating at the rorrect time, since accurate well cementing temperatures may not he known in many cases. Furthermore, in the recomple-tion of wells it is sometimes necessary to perforate cement which has set for a long period and has developed maximum or final strength. In view of the apparent need for a cementing composition having a controlled ultimate strength, an investigation was

Jan 1, 1951

By R. D. MacDonald, John Dasher, A. M. Gaudin

ION exchange is a widely used unit operation in water treatment and elsewhere in chemical industries. It has occasionally been used in hydro-metallurgy for treating plating, pickling, and rayon wastes Concerning ion exchange in extractive hydrometallurgy—for treating ore leach solutions— the literature prior to the Geneva Conference has given only hints."-" Yet most of the ore leach plants now being built or completed in the last three years use ion exchange for this purpose. Most of these plants are used for extracting uranium. The overall process for leaching low grade uranium ores and the history of its development have been given in the authors' previous papers.',' The object of this paper is to discuss certain engineering aspects of the new application of ion exchange to recover dissolved values from ore leach solutions. Necessarily the illustrations will be taken from the first application—to uranium metallurgy. Although ion exchange processes have been devised in which separation of spent residue from pregnant liquor is not required, as in the RIP process,"-'I only ion exchange operation, in which the resin is contacted with a clear solution, will be discussed in this paper. Uranium exists in acid leach liquors mainly as the cation UO," and can be adsorbed completely by cation exchange resins, but the adsorption is not selective as other cations such as Fe++, Fe+++, Al+++, Mg++, Mn", and Ca++ are also adsorbed. Uranium, however, forms complex anions with sulfate, phosphate," Carbonate"~"~ and even with chloride if the solution is strong enough.15 These anions can be adsorbed by anion exchange resins. The strong base anion exchangers have a very strong preference for these complexes and hence this adsorption is very selective. In this paper, the examples concern acid sulfate leach liquors from which uranium is adsorbed as UO,(SO,),-% and UO,(SO,),". Materials and Equipment Ion Exchange Material—Ion exchange materials include natural and synthetic; organic and inorganic; cationic and anionic; and weak, intermediate, and strong varieties. They consist of fragments or beads normally in the size range —16 to $50 mesh. The newer varieties have been developed to have greater stability, chemical strength, and/or capacity than earlier varieties. The resin used for uranium recovery is a strong base anion exchanger.* To make * Such as Rohm and Haas Amberlite IRA-400 series, British Permutit DeAcidite FF. Permutit lonac SE. and Dow Nalcite SAR. this resin, styrene plus a few percent divinyl benzene are polymerized into long chains with occasional cross links. This resin in bead form is then reacted to introduce quaternary amine groups onto most of the aromatic rings.' These resins are water-avid gels of specific gravity of about 1.1, which shrink and swell by osmosis when the ionic content of the surrounding solution is changed. The beads which are shown in Fig. 1 can be broken like an onion by rapidly replacing 15 pct H,SO, with water. This organic gel can also be broken by mechanical handling as by a centrifugal pump. Otherwise it is remarkably stable. It is unharmed by fairly strong acid or alkali, oxidizing or reducing agents, and heat up to 70°C. The strong base resin when completely in the chloride form (with a chloride anion on each quaternary amine group) holds over 107 g of C1- per dry kg. As all these chloride ions can be displaced by other anions, the resin has an exchange capacity of over 3 g equivalents per dry kg. However, the resin is sold and used by moist volume. The moist resin contains about 50 pct water, weighs about 43 lb per cu ft, has 40 pct voids, and an exchange capacity of over 1 g equivalent per liter of bed volume. Ion Exchange Equipment—Ion exchange equipment used for recovery of uranium from leach solutions is substantially identical to that developed for water treatment and other applications.'" The resin beds are held in closed tanks or columns which, in the uranium plants, have been standardized at 7 ft diam and 12 ft high. The resin bed

Jan 1, 1958

By Blake M. Caldwell, Harrold D. Owen

The continued use of permanent-type, we11 completion Iras pointed LIP the need for more powerful through-tubing perforating equipment. A new expendable shaped charge perforator has been developed in which the charges are run through the tubing in a vertical position and then opened out to a horizontal position when at the desired shooting zone. The performance of this new tool is comparable to that of conventional casing type perforators. This paper describes the new tool and its applications. Performance data in targets rind under various actual well conditions are presented. INTRODUCTION The general acceptance and increasing use of pernianent-type well completion by the oil industry has indicated the need for through-tubing perforating equipment with performance comparable to that of conventional casing type perforators.'.' To achieve the performance desired, a new, expendable, shaped charge perforator has been developed with adequate power for effectively perforating both 51/2-in. and 7-in. casing below tubing. This has been accomplished without resulting severe damage to the casing and can he obtained under extremes of pressure and temperature. The design of this new tool is a radical departure from previous concepts of tubing guns since it was felt that the two major limitations imposed by previous designs severely handicapped performance. The first of these limitations was the restriction imposed upon the length of the charges by the necessary small inside diameter of the carrier. This resulted from the necessity for compromises in designing a hermetically sealed gun small enough to allow free passage through 2-in. tubing, yet with sufficient wall thickness to withstand some minimum pressure. In an effort to overcome this limitation, the charges were in some cases placed at an angle with the axis of the well with resulting loss in effective depth of penetration. The second limitation concerned the extreme distance of the charges from the casing when the gun was fired. An example of the latter situation is demonstrated by the fact that a conventional 13/4-in. tubing gun centralized in 51/2-in., 17-lb casing has a clearance of more than 1-9/16-in. on all sides. This means that a considerable portion of the force from each charge must be expended in penetrating this 1-9/16-in. of well fluid before it even reaches the casing. With the gun centered in 7-in., 23-lb casing, the amount of fluid penetration required of each jet is in excess of 21/4-in. In a situation where the 13/4-in. gun is touching one side of the 51/2-in., 17-lb casing, the jets from the charges on the opposite side have more than 3%-in. of fluid to penetrate. A similar situation in 7-in., 23-lb casing necessitates the penetration of more than 4-9/16-in. of fluid. The penetrating effect of the charges is severely impaired in passing through so much fluid. Under such conditions, there is usually insufficient force left to penetrate the casing. Previous studies have shown the decided disadvantage of having to shoot through such an excessive quantity of well fluid.'.' The problem. then, was how to position large powerful charges close to the casing prior to firing and still manage free passage through 2-in, tubing. The solution appeared to be in some method of folding sealed charge units in an open type, expendable. 13/4-in. carrier for travel through the tubing, and unfolding them upon their arrival at the shooting zone in the casing. It was desired to control the mechanism electrically from the surface through the regular shooting cable, making the release of the charges independent of the physical characteristics of the well, yet with a provision to prevent the gun from firing until the

Jan 1, 1955

By Shiou-Chuan Sun

THIS paper presents the design and operation of a frothmeter capable of measuring the frothing characteristics of pine oils and other frothing reagents. The experimental data show that the froth-ability of pine oil is governed by: 1—rate of aeration, 2—time of aeration, 3—height of liquid column, 4—chemical composition of pine oil, 5—pH value of solution, 6—temperature of solution, and 7—concentration of pine oil in solution. The effect of mineral particles on the behavior of froth also was studied, and the results can be found in a separate paper.' The results also show that the relative froth-abilities of pine oils in the frothmeter generally correlate with those in actual flotation, provided that other factors are kept constant. In addition to pine oils, the other well-established flotation frothers were tested, and the results are included. In this paper, compressed air frothing is the frothing process performed by means of purified compressed air, whereas sucked air frothing is the frothing process accomplished by purified air sucked into the glass cylinder by a vacuum system. The term vacuum frothing denotes that froth was formed by degassing of the air-saturated liquid under a closed vacuum system. Apparatus The frothmeter, shown in Fig. 1, is capable of re-producibly measuring the volume and persistence of froth as well as the volume of air bubbles entrapped in the liquid and is capable of being used for compressed air frothing, sucked air frothing, and vacuum frothing. Fig. la shows that for compressed air frothing, the apparatus consists of an airflow regulating system, 1-3; a purifying and drying system, 4-8; a standardized flowmeter to measure the rate of airflow from zero to 500 cc per sec, 9; and a graduated glass cylinder, 13; equipped with an air regulating stopcock, 10; an air chamber, 11; and a fritted glass disk to produce froth, 12. The fritted glass disk, 5 cm in diam and 0.3 cm thick, has an average pore diameter of 85 to 145 microns. The pyrex glass cylinder has a uniform ID of 5.588 cm and an effective height of 63 cm. The inside cross-sectional area of the glass cylinder was calculated to be 24.53 sq cm, or 3.8 sq in. For sucked air frothing, Fig. lb shows that the apparatus for compressed air frothing is used again, with the following modifications: 1—compressed air and its regulating system, 1-3, are eliminated; and 2—a vacuum system, 16, equipped with a vapor trap, 15, and a vacuum manometer, 17, is added. The vacuum system can be either a water aspirator or a laboratory vacuum pump. Any desired rate of airflow can be drawn into the glass cylinder, 13, by adjusting the opening of the air regulating stopcock, 10. The sucked air stream is cleaned by the purifying and drying system, 4-8, before entering the glass cylinder, 13. When this setup is used for vacuum frothing, the air regulating stopcock is closed. The frothmeter has been used for almost 3 years and has proved to give reproducible results, as illustrated in Table I. With a magnifying glass and suitable illumination, the frothmeter also can be used to study the attachment of air bubbles to coarse mineral particles.' Experimental Procedures Except where otherwise stated, the data presented were established by means of the compressed air method. The volume and persistence of froth were recorded respectively at the end of 4 and 6 min of aeration at a constant rate of airflow of 29.3 cc per sec, which is equivalent to 71.6 cc per sq cm per min, or 462.6 cc per sq in. per min. The aqueous solution for each test, containing 1000 cc of distilled water and 19.2 ± 0.5 mg frothing reagent, was adjusted to a pH of 6.9 0.2. The volume of froth is expressed as cubic centimeter per square centimeter and is equivalent to the height of the froth column (the distance between the bottom and the meniscus of the froth). The volume of froth was obtained by multiplying the height of froth by the cross-sectional area of the glass cylinder, 24.53 sq cm. Before each test, the glass cylinder, 13, was cleaned thoroughly with jets of tap water, ethyl alcohol, tap water, cleaning solution, tap water, and finally distilled water. The cylinder with stopcock,

Jan 1, 1953

By J. J. Gilman

Zinc monocrystals, when compressed nearly parallel to their basal planes (within 20), deform in a new way; and certain deformation markings, called "-bands," are typical of this new mode. Characteristics of these bands are described. Mechanical behavior of low angle zinc monocrystals was investigated in some detail. Recrystallization did not occur upon annealing of banded specimens that were free of deformation twins. -Bands do not seem to be the result of nonbasal slip, but rather, a result of a microbuckling mechanism. u-Bands differ qualitatively and quantitatively from ortho kink bands. A dislocation model, consistent with some of the features of u-bands, is described. THE plastic deformation of zinc monocrystals with slip planes (0001) nearly parallel to the stress axis was studied in this investigation. In a previous investigation, it was shown that some low angle zinc crystals (those with slip planes making angles of more than about 2.5+1° with the compression axis), deform by ortho kinking while others (those with slip planes making angles of less than 2.5+1° with the compression axis) deform in a new way.' This new mode of deformation will be described in some detail here. Some of its striking features are: 1—it operates only for compressive stresses, 2—its temperature dependence is anomalous, and 3—it exhibits a new type of "yield-point" behavior. The mode of deformation to be discussed is not new in the sense that no previous evidence of its existence had been reported. The markings that it creates on the cleavage surfaces of zinc crystals have been described by several investigators.'" However, this is the first time that it has been made to operate under controlled conditions so that its response to mechanical variables, temperature, and aging treatments could be investigated and its geometry completely described. Apparently. Jillson' as well as Holden3 and Pratt" believed that the small bands they observed were simply small versions of ortho kink bands. In fact, two of these authors3,6 "alled the markings "micro kink bands." In spite of the precedence, it is felt that this term is not entirely suitable because these bands do not have the same genesis as kink bands, and they are qualitatively different in structure. As will be indicated by later discussion, the bands seem to be related to the tensile deformation bands observed in aluminum. The present author has chosen to call the bands "micro deformation bands," or simply ",ubands," and the deformation process, ''p-banding." Experimental Methods Crystals were grown from 99.999+ pct Zn obtained from the New Jersey Zinc Co. The details of the technique are given elsewhere." It should suffice to say that the crystals were seeded for orientation control, and precision pyrex tubes were used for molding them. About 40 crystals were used, yielding some 130 specimens. Fig. 1 illustrates the typical form of the specimens and shows how the orientation angles are defined. The crystal specimens were usually 1 3/4 in. long and about 150 mil in diameter. The angle between the slip or cleavage plane (0001) and the specimen axis is designated by xo, while a is the angle between the maximum slope of the (0001) plane and the nearest close-packed direction [1210], vary ing from 0°to 30°. The crystals were cut into specimens with a jeweler's saw and the ends were squared by filing (specimens held in V-block). Crystals of this orientation are stiff toward bending, so this procedure introduced only deformation localized at the ends. Then the specimens were polished either chemically in 50 pct HNO³ or electrolytically in 20 pct chromic acid. The mechanical tests were made with an Instron testing machine. This machine weighs by means of electric strain gages on elastic bars and is quite stiff, since the full-load deflection of its load cells is at most 0.002 in. Most of the testing was done at a

Jan 1, 1956

By Roger W. Dewey

The opposition is all kinds. There are extremists. There are quiet, sensible sounding folk who can twist numbers and facts to make their point. But they are all out to shut you down! Some of them are genuinely concerned about miners' impact on the environment. Others are just anti- society, anti-big business - small is beautiful - live naturally. The opportunity for them to make those statements on television was provided by us, the Uranium Public Affairs Task Force, as part of a media tour of the State of Idaho in May. We fielded four representatives of the industry and got 25 hours of television coverage, 22 hours of radio coverage, and print coverage by every paper in Idaho. The tour included several debates, and these clips are from two of them. Our folks creamed them! This one was so upset that he ran off the set while his mike was still plugged in, trailing studio equipment behind him. But we don't always have the opportunity to rebut them. They are making these statements all the time, everywhere they can. They have learned their trade well. They use the hearing process like A1 Hirt uses the trumpet. If the process of intervention should shut us down or prevent us from getting a license, so much to the good. But even if it doesn't - they win - for it causes delay - delay costs money and so does complying with regulation. If they can make us uneconomic, and that's not too hard to do these days, they have won. Regulation restricts the decision process. Any time the decision process IS restricted, you face the possible loss of a more economic alternative. They are out to pile every regulation on you they can and every delay they can. Initiatives! They came after us - the uranium industry - in South Dakota and Montana last year. They won in Montana. We tried to reverse it in the legislature, but they were too frightened of public reaction to do it. They did put it on the ballot for reversal in November of '82. That puts it squarely up to us to influence the public so we can win a campaign. There will be more initiatives, and at local levels as well as at state. We must join together and win! The public generally supports the continued operation of nuclear power plants. They about split on whether to build more. But they strongly support regulating the industry more stringently. Every survey reflects concern about safety and the desire for the government to take responsibility and regulate. You and I know that regulation nearly always adds cost, and only sometimes increases safety. We need to influence the creation of regulation. We need to accept responsible regulation and fight that which is counter-productive. To win in hearings, to win initiatives, to win in getting responsible regulation, we need public support. We need an informed, understanding, and supportive public. To accomplish this, we need two kinds of efforts. The first is to reach the people at the local level with local representatives of our industry. Informal conversations at church, PTA, cocktail parties, whatever. Presentations to Kiwanis Clubs, League of Women Voters, church groups - wherever we can. Facts, information in printed form, to these same local audiences with the credibility of the local sources. The second is to reach mass audiences through the media. Positive media. This can be done by advertising, but it is very expensive. We have to look to influencing the reporters and editors to get more balanced and accurate reporting. We need to get free time - interviews, debates, letters to the editors, etc. The Uranium Public Affairs Task Force was created last year to provide tools for you to use to reach these audiences (it is affiliated with the Atomic Industrial Forum). Twenty-two companies provided money and man- power. A consultant, Denver Research Group, was retained to produce materials. In this Phase I effort, we first researched what issues were of greatest concern and what were felt to be the greatest needs in materials. We determined that we did not have the funds to go out and do the job for the industry, so we decided to develop tools for the Industry to go out and do the job for itself. From the research, we determined what tools we should develop for you to use. We first developed a set of the quest- ions most likely to be asked of you and the issues most likely to be thrown up to you. We have developed a loose-leaf notebook. Each page contains one of those questions or issues, a short verbatim response that you can use, a short discussion of the subject, and references you can cite or research for further information. It is organized by subject: tailings, water radiation, etc. This book is an extremely handy tool for anyone in the industry. Each uranium location should have at least one.

Jan 1, 1982

By D. R. Wieland, A. N. Barron, A. R. Hendrickson

A definite relationship has been found between the reactivity of flowing hydrochloric acid and its shear rate in a carbonate fracture. Both flow velocity and fracture width affect the acid reaction rate. Laboratory studies were conducted on acid reactivity at different flow velocities through horizontal-linear fractures, using 15 per cent hydrochloric acid at 80' F and approximately 1,100 psi. Fracture width varied from 0.02 to 0.20 in. These data provide a new basis from which the spending time and penetration of the acid can be estimated. Equations were derived expressing the relationship between injection rate, fracture width, acid concentration, time and fracture height, for linear and radial fracture systems. Because the penetration of the acid before spending is closely related to the extent of productivity increase resulting from an acidizing treatment, these data provide a valuable insight into some of the controlling factors that must be taken into consideration during treatment preplanning. INTRODUCTION Acidizing of carbonate reservoirs to improve production characteristics has been widely practiced since 1932. Originally, it was assumed that the acid uniformly penetrated natural formation pores and flow channels, enlarging them and thereby increasing their flow capacity. Little consideration was given to the reaction rate of the acid, or how far it would penetrate away from the wellbore into the formation, before spending. It has been shown recently' that, unless fractures are present in the rock, very little penetration is attained before spending, and the benefits of the acidizing treatment are largely confined to the immediate vicinity of the well-bore. Therefore, acid treatments may be classified into two categories: (1) matrix acidizing, in which the acid flows through multiple: formation pores; and (2) fracture acidizing, in which the bulk of the acid travels through fractures in the rock, whether natural or induced. When acidizing treatments are conducted at pressures of sufficient magnitude to open and extend such fractures, it is often desirable to inject a propping agent to hold the fracture open after the treating pressure has been released, thus providing a highly conductive flow channel through the rock. In some cases where acid attack produces surface irregularities, the resulting flow passages can be sufficient to provide high conductivity without use of a propping agent. During injection, the acid dissolves the carbonate rock with which it comes in contact until it is spent. Deeper penetration of the spent acid into the formation produces no appreciable benefit (if no propping agent is used) because the unetched fracture faces will rejoin when treating pressures are released and the fracture will "heal", with negligible resultant conductivity. The spending time of the acid thus becomes important in determining how far from the wellbore the improved-conductivity zone extends. The spending time of acid after injection into carbonate rock depends on the rate at which the acid reacts with the rock. This in turn is controlled by a number of factors, as previously reported. 2-4 These include temperature, pressure, acid concentration, rock composition, injection rate and the area-volume relationship between the acid solution volume and the surface area of the formation flow channels through which it penetrates. Thus, in matrix acidizing, where an extremely large area is exposed to the acid, spending is rapid. In contrast, spending time is prolonged in open fractures, where the area-volume ratio is much lower. Thus, greatest penetration of the acid before spending will be achieved in acid-fracturing treatments where fractures are held open by hydraulic pressure. Perkins and Kern5 have presented equations for fracture-width determinations which, for hard formations such as limestone or dolomite, indicate that high injection rates and viscous fluids are required to create wide fractures. Increasing the injection rate will, in itself, produce faster reaction rates. Because this type of reaction can be considered to be first-order diffusion-controlled, '-' the rate of movement of fluid past the rock surface affects the thickness of the diffusion layer, in turn affecting the reaction rate. Staudt, et al,10 report the results of tests in which a rotating marble cylinder was immersed in acid, and the weight loss at different speeds determined. Such test conditions, however, are not truly indicative of the situation in a formation during an acid-fracturing treatment. This paper reports reaction-rate data obtained under

By R. B. Full

DURING the summer of 1949, the United Nations Scientific Conference on the Conservation and Utilization of Resources met at Lake Success. As summed up by one writer, the purpose was: "That everyone will try to learn how to feed more people from an acre of land." That objective is a stimulus to the phosphate rock industry as well as other industries. Phosphate rock was mined first in the Canadian province of Quebec where it was found as an apatite in pockets. In 1863, mining began near the Rideau Canal, first in open trenches and quarries until shafts were sunk. These operations were difficult since hard rock had to be drilled and blasted. Steam was the source of power. The phosphate rock was hand-selected and recovery ranged from 6 to 10 pct, averaging 70 pct to 85 pct bpl (bone phosphate of lime). Costs, f.O.b. Montreal, were $14 a ton, and selling prices averaged $17 a ton. In 1885, about 29,-000 tons were sold on this basis, but by 1892 the output fell to 8000 tons at an average price of $15 a ton at Montreal. South Carolina Deposits Phosphate rock was first discovered in South Carolina in 1837, but a mining company was not formed until 1867. It was immediately successful, and the industry continued to develop and ship an increased tonnage of land and river rock, shipments increasing from approximately 20,000 tons in 1868 to 537,000 tons in 1890, when the shipments began to decrease. In 1892 only 350,000 tons were shipped. In this field there were two types of deposits: land rock and river rock. More land rock tonnage was mined until about 1887 when river rock production began to equal it. The land rock was mined adjacent to and in the river marshes by pick, shovel, and wheelbarrow. Each worker dug daily a space 6 ft wide, 15 ft long and 6 ft deep, which was overburden, and then loaded about 3 cu yd of phosphate rock into wheelbarrows. His helper would wheel the barrows to small steam railroads up to 250 ft distant. Water was kept from the pits by steam pumps connected to railway cars. The phosphate rock was loaded into railroad cars, hauled and dumped into a log washer to which water was added, and run over trommel screens to eliminate the clay and loose sand. Steam shovels for loading were introduced about 1891. River rock was mined by floating dipper dredges and at low tide was loaded into barges by laborers using oyster tongs, the large lumps being hand-loaded. Barges and bateaus were unloaded by wheelbarrows which were wheeled to the same type of land rock washers. The washer product was piled with stacked pine cordwood in alternate layers and fired for drying. Usually this was done under sheds along the river bank. The dried rock then was transferred to ships, usually three or four-masted schooners. Cost of production of both land and river rock, including $1 a ton royalty paid to the state for state-owned property, was approximately $4.25 a ton averaging 60 pct bpl. The selling price was approximately $7 a ton f.O.b. In 1886 the courts handed down a decision permitting anyone to mine, and many additional companies were organized immediately. As a result, the selling price was lowered to a profit margin of 50Ø or less per ton. The South Carolina mines were eventually shut down when a more profitable source of phosphate rock was discovered in Florida. Florida Deposits Phosphate rock was discovered in Florida in 1882, and in 1883 a small quarry was opened near Hawthorne. In 1886 phosphate rock of good grade was discovered on Peace River, and the Arcadia Phosphate Co. was formed. In 1888 the company began shipments amounting to 3000 tons that year. During the latter part of 1888, hard-rock phosphate was discovered near Dunnellon. Exploration extended rapidly, and deposits of hard-rock phosphate were mined over a wide area so that shortly after 1900

Jan 1, 1952

By H. R. Froning, R. O. Leach

In wertability alteration flooding, a chemical agent is rnoved through a reservoir by the flood water to increase oil recovery by decreasing the degree of wetting of the rock by the oil. Substantial amounts of the chemical may be lost during movement through the reservoir. The extent of the loss, and therefore the economics of the process, depends in some cases on factors which are difficult to reproduce in the laboratory. Therefore, a short-duration, low-cost field test method is needed to permit evaluation of chemical requirements under actual field conditions. This paper describes a small scale rest conducted at a single well for measuring chemical requirements, thereby giving a more reliable evaluation of this important factor in the applicability and economics of the process. In the rest a small wafer slug containing the chemical agent and a nonadsorbed tracer is displaced into the reservoir by a known volume of wafer. The well is then placed on producrion. Chemical loss per barrel of pore volume contacted is calculated from [he fractional recoveries of the agent tested and the nonadsorbed tracer. The method has been used to determine within the actual reservoirs the chemical requirements for both a sandstone and a dolomite reservoir. Several chemical agents are potentially available for wettability alteration flooding, although none is universally applicable. For some applications of the method, chemical costs per barrel of additional oil recovered can be substantially less than one dollar. INTRODUCTION Wettability alteration flooding provides a means of increasing oil recovery from reservoirs by decreasing the degree of wetting of the rock by the oil and increasing the displacement efficiency of the flood water. Earlier studies demonstrated a relationship between oil recovery during waterflooding and the degree of wetting of a rock surface by an oil. The application of wettability alteration flooding to the Harrisburg field of Nebraska provided a field test' of this recovery process. Subsequently, additional laboratory and field tests have developed additional procedures for evaluating wettability alteration flooding, and have indicated where the process may be applicable. Applicability of this process to specific reservoirs is determined by a progression of tests to determine sus- ceptibility of the reservoir to alteration of its wettability, to indicate the degree of recovery improvement and to estimate the amount of chemical required to process the reservoir. The economics of applying improved oil recovery processes depends not only upon the degree of improvement in oil recovery achievable by the process but also upon the process costs and the timing of the income and the investment. Emphasis in this paper is on the expenditure aspects of the process. The work reported in this paper indicates that the chemical investments required for wettability alteration flooding are substantial. For evaluating the economics of a potential flooding application it is imperative that a sound estimate of the chemical requirements be made for the reservoir. Generally, true reservoir conditions are not adequately simulated in laboratory chemical propagation tests. Because of wide well spacings, many years might be required to obtain chemical propagation data from conventional pilots or inter-well tests. Consequently, n short-duration, low-cost method is needed to determine chemical requirements in the field. The potential applicability of wettability alteration flooding is discussed, as well as the economics of wettability alteration with respect to the inherent and imposed restrictions on the timing of income and investments. DETERMINATION OF CHEMICAL REQUIREMENTS In the process of moving a chemical bank through reservoir rock, some of the chemical agent lags too far behind the flood front to be effective or is otherwise lost to the reservoir system. The extent to which these losses occur overshadows the reductions in chemical concentration due to diffusion and to mixing with the reservoir fluids. Experience indicates that almost without exception, chemicals which induce a wetting change undergo either sorption reactions or chemical reactions with mineral constituents of the pore surfaces. Other reactions may occur between the added chemical and the reservoir oil and water. Even in limiting consideration to reactions of the relatively inexpensive inorganic salts, bases and acids, the reactions may be exceedingly complex. Reservoir pore surfaces consist of more than silica in sandstone reservoirs, and more than calcite or dolomite in carbonate reservoirs. Many mineral species are present, each exhibiting specific tendencies to react with an injected chemical. The reactions which occur can consume enough of the agent to have an important effect on economics. These reactions can cause a change in pH of the chemical bank, or may remove some of the active chemical by precipitation, adsorption or reaction to form a new chemical which may or may not be effective in

By E. T. Turkdogan, Klaus Schwerdtfeger

OelSEN and Maetz1 detected some 20 years ago the existence of a miscibility gap in iron oxide-CaO-P2O5 slags melted in iron crucibles at about 1400°C. Because of the importance of this system for the dephos-phorization of steel in the basic Bessemer process, equilibria between liquid iron and selected iron oxide-CaO-P2Q slags have been measured since by numerous investigators.2-5 When in equilibrium with metallic iron, the iron oxide of the slag is present mainly as FeO. In connection with oxygen-blowing steelmaking processes, it is useful to know the phase relations in the slag system at higher oxygen pressure, when major parts of the iron oxide are present as Fe2O3. This problem was investigated by Turkdogan and Bills7 by equilibrating the oxide mixtures contained in platinum crucibles with CO2-CO mixtures at 1550°C. It was found that increasing the Fe2O3 content decreases the composition range of the miscibility gap strongly so that the miscibility gap has almost disappeared at pco2/pco = 75. This result was refuted by the careful work of Olette et a1.,''' who equilibrated their slags with controlled Ha-H2-Ar gas mixtures. Their equilibrium measurements, at 1600°C and at oxygen pressures of 5 x 10"* and 10"5 atm, showed that the oxidation state of the iron has almost no influence on the formation of the miscibility gap. The present experiments were undertaken to check the previous results of Turkdogan and Bills. The experiments were performed at 1625°C in the strongly oxidizing atmosphere of air (PO2 = 0.20 atm) for which no experimental data are available. About 10 g of slag were melted in platinum crucibles and held at constant temperature for 1 hr. After equilibration, the crucible was rapidly pulled out of the furnace and cooled in air. The platinum crucible was removed from the sample. The two slag layers were carefully separated with a small diamond disc, and the surface of the top layer, which may have changed its oxidation state during cooling, was removed. The slags were crushed and analyzed chemically for CaO, P2O5, Fe2+, and Fetotal. The starting mixtures were prepared by sintering the desired amounts of reagent-grade 2CaO . P2O5 - H2O, CaCO3, and Fe2O3. Sintering and subsequent crushing were done three times to ensure homogenization. Molybdenum wire resistance heating was used. The furnace was provided with a recrystallized alumina reaction tube which was left open to air at the top. The temperature was controlled electronically. The reported temperature was measured with a Pt/Pt-10 pct Rh thermocouple and is estimated to be accurate within +5°C. The composition of the equilibrated melts is given in Table I. For the graphical illustration of these quaternary slags the type of projection suggested by Trömel and Fritze10 was used. In this representation, Fig. 1, the composition point of a mixture within the tetrahedron Fe2O3-CaO-P2O5-FeO is projected into the Fe2O3-CaO-P2O5, triangle (triangle I) so that the direction of projection is parallel to the side FeO-Fe2O3, and into the triangle Fe2O3-P2O,-Fe0 (triangle 11) so that the direction of projection is parallel to the side CaO-P2O5, of the tetrahedron. The projected point has the coordinates wt pct CaO, wt pct P205, and wt pct (FeO + Fe2O3) in triangle I and wt pct FeO, wt pct Fe2O3, and wt pct (CaO + PzO5) in triangle 11. Both triangles are turned into the same plane around the Fe203-P20, side of the tetrahedron. An illustration of the projection of a quaternary point in the present system is shown in Fig. 1. The advantage of this type of projection is that all four components for an equilibrium curve can be read directly from the diagram. The present results are shown graphically in Fig. 2. The curves depicting the miscibility gap are dashed in parts where no experimental points were obtained. The composition range covered by the miscibility gap

Jan 1, 1968