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By E. Miller, J. M. Eldridge, K. L. Komarek

The thermodynamic properties of liquid Mg-Ge alloys have been determined between 1000°and 1500°K by an isopiestic method. Germanium specimens, heated in a temperature gradient and contained in covered graphite crucibles of special geometry, were equilibrated with magrtesium vapor in closed titanium tubes. The crucible design allowed free access of magnesium vapor to the samples during the equilibration to form alloys of magnesium and germanium, but prevented magnesium losses from the crucibles on quenching the titaniuin tubes to terminate the experimental runs, thus preserving the equilibrium alloy compositions. The activities and partial molar enthalpies of magnesium and the integral thermodynamic properties of the system were calculated from the experimental data. THE Mg-Ge phase diagram' shows one congruent melting compound, Mg2Ge, of essentially stoichio-metric composition, two eutectics, and very limited terminal solid solubilities. Very little information is available on the thermodynamic properties of the Mg-Ge system. The free energy of formation of Mg,Ge was recently deter-mined2 by a Knudsen cell technique in the temperature range 610° to 760°C. The standard enthalpy of formation of Mg,Ge was measured calorimetrically by Bever and coworkers.3 The present study was undertaken as part of a general investigation of the thermodynamic properties of the homologous series of Mg-Group IVB systems, i.e., Mg-Pb,4 Mg-Sn,5 Mg-Ge, and Mg-Si. An isopiestic technique was used which was developed by the authors5 for investigating the thermodynamic properties of liquid Mg-Sn alloys. Specimens of the nonvolatile component, contained in covered graphite crucibles, are heated in a temperature gradient in an evacuated and sealed titanium reaction tube, and equilibrated with magnesium vapor of known pressure. The method employs crucibles of special geometry which preserve the high-temperature equilibrium composition of liquid alloys having a highly volatile component such as magnesium on termination of the experimental runs by quenching the crucibles to room temperature. EXPERIMENTAL PROCEDURE First reduction germanium of 99.999+ pct purity (Eagle-Pitcher Co., Cincinnati, Ohio) and 99.99+ pct magnesium metal (Dominion Magnesium Ltd., Toronto, Canada) were used. The graphite crucibles were machined from high-density (1.92 g per cu cm) graphite rods (Basic Carbon Corp., Sanborn, N.Y.) which had a maximum ash content of less than 0.04 pct. The non-reactivity of graphite with germanium at the temperatures used in this study had been previously established by Scace and Sleck.6 The experimental procedure has been previously described in detail.5 The selection of a particular crucible geometry for a run was determined by a combination of imposed experimental conditions, the principle being that more tightly covered crucibles were required to preserve alloy compositions during quenching when higher magnesium pressures and higher specimen temperatures were used. Depending upon the composition range of the equilibrated alloys the source of the magnesium vapor was either pure magnesium or a two-phase mixture of Mg2Ge + Ge-rich liquid of known magnesium pressure. The experimental runs can be divided into the following three groups on the basis of crucible geometry and magnesium source material. Crucibles with Small Holes and Pure Magnesium Reservoirs. The crucible dimensions were identical to those of the Mg-Sn investigation5 except that the hole diameters were reduced to 0.010 in. because of the higher temperatures and higher magnesium pressures involved in the Mg-Ge system. During an equilibration run, magnesium vapor diffused from the reservoir to each specimen through the small holes, one drilled through the crucible lid and two others drilled through graphite baffles positioned vertically inside the crucible between the lid hole and the specimen. Since the magnesium pressure was high, i.e., in the range 117 to 277 Torr, during the equilibration time of approximately 24 hr, equilibration was not impeded by these holes. A specimen composition at equilibrium was fixed by the relative temperatures of the specimen and the reservoir, and by the thermodynamic properties of the system. Upon brine quenching the titanium reaction tube to end a run the vapor pressure of magnesium above the liquid alloys decreased exponentially with decreasing temperature, and the small cross-sectional areas of the holes (4.9 x 10"* sq cm) drastically reduced magnesium losses from the crucibles. Because of its low vapor pressure, germanium losses from crucibles during a run were at most 0.2 mg for pure germanium and correspondingly less for the alloys. This crucible geometry satisfactorily retained the equilibrium alloy compositions on quenching for magnesium-rich (from 3 to 33 at. pct Ge) alloys provided their temperatures were below the melting

Jan 1, 1967

By G. L. Fisher, R. H. Geils

The effect of small amounts of columbium (<0.01 to 0.10 pct) on the ?-a transformation occurring during the continuous cooling of a low carbon Ni-Cu steel was investigated. Dilatometer specimens were aus-tenitized at 950" and 1068?C and cooled at 17? and 375 C° per min. Columbium caused a marked depression in the ?-a transformation temperature except when cooling at the slower rate from 950°C. The effec of columbium on the transformation temperature was greater the higher the austenitizing temperature and rate of cooling. A maximum depression of 92 C" was observed. Metallographic examination of specimens of <0.01 and 0.07pct Cb steels heated at 1200°C for 1 hr and cooled at various rates showed that columbium had a major effect on the ferrite morphology. The fer rite in the columbium -free steel remained equiaxed at cooling rates as high as 440 C? per min while the columbium-bearing steel exhibited mixed structures o equiaxed and bainitic ferrite at cooling rates as low as 130 C° per min. The ? grain boundaries in the columbium -free steel provided the ferrite nucleation sites in rapidly cooled specimens. There was a complete absence of nucleation at these sites in the colum bium-bearing steel. It is concluded that columbium depresses the transformation temperature by suppressing ferrite nucleation at the austenite grain bound-aries. In this respect the effects of columbium are analogous to those of boron in low C-Mo steels. It is well known that small columbium additions can substantially strengthen plain carbon steels. As little as 0.02 pct Cb can increase the yield strength of mild steels by 10,000 psi.1 A fine precipitate of CbC has been observed in columbium-bearing steels2 and is generally thought to be responsible for the strengthening. Little attention has been devoted to the effect of columbium on the ?-a transformation. Webster and woodhead3 have studied the effect of columbium on the isothermal proeutectoid ferrite reaction in mild steels. They found similar transformation behavior in steels both with and without columbium additions. However, as the austenitizing temperature increased, the incubation time for the start of the ferrite transformation became longer in the columbium-containing steel. Morrison1 found that the addition of 0.03 pct Cb to a C-Mn steel lowered the transformation temperature by 50 C° during cooling from 1200°C at a rate of 80 C" per min. The strengthening effect of columbium has recently been utilized in an age-hardenable, low-alloy steel containing copper and nickel.4 A small amount of columbium has a substantial effect on the as-rolled strength of this steel. By increasing the columbium level from <0.01 to 0.13 pct the as-rolled yield strength is increased by 15,000 psi. Columbium also significantly lowers the ?-to-a transformation temperature of this steel during continuous cooling from the austenitizing temperature. Because of the low carbon level in this steel (0.05 pct max), it is almost entirely ferritic. Thus, it offers the opportunity of studying the effect of small columbium additions on the proeutectoid ferrite reaction. Of particular interest in this study was the reason for the marked lowering of the transformation temperature by columbium during continuous cooling. EXPERIMENTAL PROCEDURE Materials. The compositions of the steels used in this investigation are shown in Table I. The steels were 30-lb air induction melts. They were forged to 4 by 8 by 1 in. plate at 1230°C, air cooled, and then reheated to 1230°C and cross-rolled in two passes to in. plates. Dilatometry. A Leitz Bollenrath dilatometer was used to record the transformation during continuous cooling from two different austenitizing temperatures. The dilatometer specimens were + in. in diam and 2 in. long. Oxidation and decarburization of the specimens was prevented by maintaining a small positive pressure of dry argon in the dilatometer furnace and by plating the specimens with 1 mil of Cu. For the lowest cooling rate, 17 C" per min, the temperature of the specimen was measured with a Pt-Pt 10 pct Rh thermocouple placed in a & in. diam well in the center of the specimen. During air cooling, 375 C° per min, this method of measuring the temperature interfered with the operation of the dilatometer. However, it was found that the temperature of the specimen could be measured accurately by placing a thermocouple in an identical specimen in a holder adjacent to the one being used to operate the dilatometer mechanism. The dilation-temperature curves were recorded on photographic film and then converted to volume percent ferrite-vs-temper-ature curves. The cooling rates obtained with the dilatometer are shown in Table 11. Cooling rates of

Jan 1, 1970

By N. Mungan

Formation damage, i.e.. reduclion in permeability, has been generally attribuled to clay minerals which expand or disperse upon contact with water that is less saline than the connate water. Luboratory, studies show that penneahility reduction can also occur in formalions containing only nonexpandable clays such as illite or kaolinite, and can be caused also by changes in pH. Furthermore, pH changes can damage even formations that are essentially free of clays. It is suggested that permeability reduction is due to the small passages being blocked by particles, which may be dispersed clays, cemenlion material or other fine parricles. These particles are dislodged by dispersion of clays due to changes in salinity or by dissolution of calcareous cement by acids, or of silicaceous cement by alkaline solutions. In working with reservoir cores, it was found that extracted cores damaged more easily and extensively than nonextracted cores. The extent of damage depended also on tenlperatltre. INTRODUCTION Permeability is an important property of porous media and has been the subject of many studies by engineers and geologists. Many of these studies are conccrned with formation damage, i.e., reduction in permeability, resulting from exposure of oil-producing formations to water substantially less saline than the connate water. This effect causes understandable concern since during drilling, completion and production phases formations are often exposed to fresh water. The damage resulting from contact with relatively fresh water has been attributed to expansion and dispersion oF clay minerals. During laboratory investigation of the use of NaOH as a wettability reversal agent to increase oil recovery from oil-wet reservoirs, several cores used in the displacement studies suffered loss in permeability. Despite the traditional usage of NaOH for conditioning aqueous mud systems, the role of the caustic filtrate in wellbore damage seems to have been overlooked. Browning2 as recently reported on the effects of NaOH in dispersing clay minerals but he was concerned only with complications that may arise in drilling massive shale beds. The following study was made to examine the role of pH and salinity changes in core damage. Where cores from reservoirs were used, tests were performed with extracted and nonextracted cores both at room and reser- voir temperatures, since it was felt that the test environment and core condition may affect the results. Because of its limited coverage and exploratory nature, this study is not intended to provide answers to field formation damage problems. It is hoped that it will encourage research into new aspects of the permeability reduction problems, particularly those allied to new recovery and production processes. PROCEDURE In all permeability tests, fluids were pumped through the cores at a constant volumetric rate. Only deaerated fluids and reagent grade chemicals were used. The fluids were passed through two ultrafine filters before injection to remove any entrained particles. The cores, with the exception of the unconsolidated cores, were mounted in Hassler holders. Water was used to transmit pressure to the sleeve. The inlet endpiece had two entry ports which permitted scavenging one fluid with another to avoid any mixing in the small holdup volume. The cores were flushed with CO2 gas, evacuated for 5 to 6 hours and saturated with the first liquid at a pressure of 1,000 psi for 24 hours to eliminate any free gas from the cores. Pressure differences up to 20 psi were measured by transducers, calibrated in inches of water and continuously recorded. For greater pressure drops, gauges were used. All reservoir cores were cleaned with a light refined mineral oil, then with heptane, and finally dried with CO2. Compatibility tests showed that no precipitates formed when mineral oil and the crudes were mixed. Some cores were extracted in Dean-Stark-type solvent extractors using xylene and trichloroethane and dried in a vacuum oven at 450F. Each test consisted of a sequence of water, test solution, and again water flow. RESULTS AND DISCUSSION STUDIES IN BEREA CORES Salinity Contrast Berea cores 2-in. in diameter and 12-in. long were cut from sandstone quarried in Cleveland, Ohio. The clay minerals were identified by X-ray diffraction to be chlorite, kaolinite, illite and incerlayered illite. Flow of fresh water or 30,000 ppm brinc does not cause any permeability reduction (Fig. 1). However, after injection of brine the core is readily damaged by fresh water. Damage starts almost instantly as the fresh water injection is begun, and at a cumulative injection of 1.2 PV fresh water, the permeability has dropped from 190 to 0.9 md. Upon continued injection, the effluent contains clay minerals dislodged from the core. The final core per-

Jan 1, 1966

By C. B. Alcock, L. I. Cheng

By the use of radiochemical methods for the study of the gas-liquid equilibria at low temperature, and for the determination of the sulfur contents of metal beads which had been equilibrated with H2S/H2 mixtures of known sulfur potential, it has been possible to obtain the liquid solubility and the free energy of solution of sulfur in liquid tin and lead at temperatures between 500°and 680°C. THE gas-liquid equilibrium method has proved in the past to be most successful in the determination of the thermodynamic behavior of dilute solutions of sulfur in liquid metals.1,2 One of the basic requirements for success with this method is that the volatility of both the metal and its lowest sulfide should be small, otherwise sulfide will be deposited at the cool end of the furnace, where it may react with the outgoing gases to form either sulfur-rich lowest sulfide or higher sulfides. The resultant value of the apparent equilibrium constant will then be lower than the correct one. This argument applies even at sulfur potentials below that in equilibrium with a separate condensed phase of the lowest sulfide at the reaction temperature, T. The mass of sulfide which is deposited at the cold end of the furnace, and hence the extent to which further reaction occurs with the outgoing gases, depends on the time taken for equilibrium to be reached between metal and gas. Since this will depend principally on the bulk of the metal phase which is used, one should clearly attempt to uie as small metal samples as possible. These considerations are important in the study of dilute solutions of sulfur dissolved in liquid tin and lead which both have moderately high vapor pressures as metals and form volatile sulfides. The limit on the size of the metal samples which may be used is set chiefly by the difficulties of analysis for very small amounts of sulfur. The oxygen or carbon dioxide combustion method, followed by iodimetric determination of the sulfur dioxide which is formed,has been found to be successful for the determination of small amounts of sulfur in copper, iron, cobalt and nickel.4 This method was unsatisfactory for sulfur dissolved in tin and lead, mainly because the sulfur dioxide was to some extent absorbed by the copious tin or lead oxide deposits which were formed on the walls of the combustion tube. Furthermore some of the sulfur was found to segregate on the surface of the beads as flaky sulfide crystals which would easily be lost in the transfer of a bead from a boat in the gas equilibration apparatus to one in the combustion apparatus. Oxidation in aqueous media to sulfate ion followed by precipitation as barium sulfate was, therefore, adopted as the analytical procedure. The gas-metal equilibrium experiments were all carried out with radioactive sulfur and thus the analysis involved the counting of barium radiosulfate. Furthermore the use of the radioisotope meant that the approach to the gas-metal equilibrium could be followed continuously by gas counting.&apos; The metal beads were held separately in glass crucibles during equilibration and were transferred from the furnace to the beaker for dissolution in nitric acid still in the crucibles, and thus the possibility of sulfur loss by detachment of the sulfide segregates was eliminated. The temperature range of this investigation was 500° to 680°C. EXPERIMENTAL APPARATUS AND METHOD The apparatus consisted of two furnaces placed in series in a gas recirculation system, Fig. 1. One furnace F1, which was vertical was used to heat the alumina crucible, A, holding six metal beads in separate glass crucibles. The beads weighed between 300 and 700 mg each. The crucible assembly was introduced and removed from the furnace mechanically under a stream of oxygen-free argon. The other furnace, F2, was horizontal and was used to heat a cobalt Co9S8 mixture, held in an alumina boat, and made with radiosulfur containing about 1/2 millicurie per g of sulphur. This mixture, which was finely powdered, was used as a source of known H2S/H2 mixtures6 for a given furnace temperature. The recirculation system also contained a gas re-circulation pump (P), an end window Geiger-Miiller counter (N)—placed downstream of F1 so as to monitor the H2S pressure in the gas leaving this furnace— a sample volume for chemical analysis of the gas phase (G), gas drying tubes (D), filling taps and other standard ancillary equipment. The gas sampling volume was principally used in the cali-

Jan 1, 1962

By R. F. Lynch, J. D. Wood

The Al-1.0 pct Mg alloy 565 7 was studied using optical microscopy and electron microprobe X-ray analysis. Constituent particles were found to exist inter-dendritically in the as-cast material in a region of precipitate free a -aluminum. Five phases besides a fine precipitate and a-Al were identified in the cast structure: Fel3, Fe2Al7, Mg2Al3, CUMgAl2, and Cu2FeAl7. Thermal treatments conducted for 100 hr at 1180°, 1130°, 1080°, 1030°, 980°, and 880° F revealed a general dissolution and spheroidization of the in-terdendritic constituent network observed in the cast structure. The principal constituents present in the thermally treated structures were FeAl3 and Fe2Al7 with the relative amount of Fe2A17 to FeAl3 increasing with a decrease in the treatment temperature. The phases Present in the wrought structure were identical to those observed after the thermal treatments, with the constituent particles strung out in the direction of rolling. ALLOY 5657 is a nonheat-treatable commercial purity Al-1.0 pct Mg alloy utilized extensively because of its bright finishing characteristics. This investigation was conducted to determine the constituents present in 5657 alloy, and to study the effect of extended thermal treatments on morphology. Numerous studies have been carried out to establish the equilibrium diagrams for various aluminum systems,1-3 with phase identification based on X-ray analysis, morphology, and the etching response of relatively large particles. Phragmen4 conducted a study of the phases in aluminum eutectic systems and compiled a "corrected" table of etching responses, drawing on his work plus that of Schrader,5 Keller and Wilcox,6 and Mondolfo.7 A review of the original work of Keller and Wilcox, and Mondolfo, which was concerned with the constituents found in commercial alloys, reveals that in numerous cases their etching responses differ from those reported by Phragmen and from each other. These inconsistencies may occur because a specific constituent will react to a given etch in a varying manner depending upon its size, the elements dissolved in the phase, the other constituents surrounding a phase, and the solute content of the matrix. Work with a commercial orientation was conducted on alloys 2024 and 3003 by Sperry8,9 and on alloy 3003 by Barker,10 where the relationship between the phase diagram and the nonequilibrium structure of an alloy was examined. Backerud11 investigated the A1-Fe binary system and found that at high cooling rates the equilibrium eutectic reaction forming a-A1 and FeA13 is replaced by another lower temperature eutectic reaction forming a-A1 and metastable FeAl6, a constituent first identified by Hollingsworth et al.12 Most of the above mentioned studies were conducted on materials having a significantly greater alloy content than 5657 alloy, where the relatively small size and sparse distribution of second phase particles hinders the process of identifying constituents. EXPERIMENTAL PROCEDURE Material with a composition as given in Table I was examined in the as-direct chill cast, hot rolled and cold rolled conditions, and after thermal treatment of the cast structure. Thermal treatments were terminated by a water quench. Microscopic examination was conducted under various lighting conditions following the application of standard etchants as specified in Table 11. A semi-quantitative electron microprobe X-ray analysis was conducted for Al, Mg, Fe, Cu, Si, Zn, and Ti. RESULTS AND DISCUSSION Microstructure of the As-Cast Material. Particles of second phase material were found to exist inter-dendritically, principally in regions of precipitate free a-Al, as illustrated in Fig. 1. Adjacent to the ingot edge was a region of inverse segregation, resulting in an increased amount of second phase material containing large sized particles which aided in phase identification. Phase Identification. Cast Structure. Five phases besides a-Al and a fine precipitate were identified using optical microscopy and electron microprobe X-ray analysis, as presented in Tables II and III, respectively. FeA13 and Fe2A17 are often found with Fe2A17 forming a sheath around the core of FeAl3, resulting from an incomplete peritectoid reaction. These phases have a nearly identical appearance under white light, although they are easily differentiated under crossed polarizers, as characteristically illustrated in Figs. 2(a) and 2(b), respectively. Microprobe analysis con-

Jan 1, 1970

By Thomas C. Wilder, Walter E. Galin

Measurements of the electromotive force of the cell at 995°C have shown that the cell may be used to detennine the zinc content of molten Cu-Zn alloys to the nearest 0.05 wt pct. The cell is used for brass melts to which no ZnO is added intentionally, because essentially all oxygen in the Cu-Zn-O system is present in the form of ZnO. The cell is also used to detev,nine the thernodynamic activity of zinc for Cu-Zn al1oys of the brass composition range, from which the equilibrium partial pressure of zinc for such alloy may also be calculated. The electrochemical measurement of the concentration of the most active metal in other engneering- alloy systems by a similar technique is also considered. IN recent years there have been many studies concerning the direct electrochemical measurement of oxygen content in molten engineering metals.&apos; These investigations employ galvanic cells in which the electrolytes are various solid mixed-oxides which conduct current by movement of oxide ion vacancies under the influence of an oxygen potential gradient. Similar cells have also been used to measure the thermodynamic properties of mixing of solid alloys,&apos; one component of which has a much greater affinity for oxygen than the other component(s). One alloy system of obvious engineering importance comprised of two metals of greatly differing affinity for oxygen is copper-zinc. At the casting temperature of molten brasses, the zinc concentration is very difficult to control not only because its vapor pressure is very high, but also because of its high affinity for oxygen to form ZnO. Inasmuch as brass castings may be required to have a zinc concentration within very narrow limits, it would be advantageous for the brass industry to have a means for quickly measuring the zinc content in the molten alloy just prior to pouring. A calculation based on earlier thermodynamic studies of oxygen and zinc in dilute solution in binary Cu-O and CU-Z&apos; alloys indicates that the oxygen concentration of molten 70 wt pct Cu-30 wt pct Zn brass must be much less than 1 ppm for the separate phase ZnO not to exist in the ternary Cu-Zn-O system. Thus it can be assumed that all molten brasses, both in the laboratory and in the foundry, are sufficiently saturated with oxygen for the separate phase ZnO to be present. In view of the foregoing reasons, the cell investigated at 995°C for the purpose of measuring the zinc concentration in molten copper-zinc alloys, particularly those of brass compositions. In this cell the left hand electrode is a reference electrode comprised of a mixture of the pure powders of nickel and nickel oxide, to which a platinum contact is made. The electrode of interest is a molten Cu-Zn alloy of unknown zinc concentration with a tantalum wire contact. The electrolyte is the commercially available (Zircoa) calcia stabilized zirconia. The ZnO in the Cu-Zn electrode compartment is not intentionally added, but is naturally present as a result of the infinitesimal amount of oxygen required for its formation in this alloy system. In cell [A] nickel oxide is reduced at the cathode where O= represents the electrolyte. At the anode dissolved zinc of the Cu-Zn alloy is oxidized to form ZnO(s) in its standard state Thus the overall reaction is for which the molar free energy change is where 5 is the Faraday equivalent (23,063 cal per v equivalent), is the electromotive force of the cell in volts after correction for the Pt-Ta thermocouple, FOf (ZO) and AFF (NO) are the standard molar free energies of formation of the respective oxides, and a2n is the thermodynamic activity of zinc in the molten Cu-Zn alloy. The reference state of zinc in this case is taken to be the pure liquid. At constant temperature all the terms of Eq. [I] are constant except and azn Thus since all other participants in the electrochemical reaction are in their standard states, the change in the electromotive force of Cell [A] is represented only by a change of the thermodynamic activity of zinc at constant temperature. The zinc concentration is related to the activity by sampling the alloy for chemical analysis after cell measurements are taken. EXPERIMENTAL A typical cell construction is shown in Fig. 1. This and other experimental details were similar to that described earlier with the exception of some minor modifications which are detailed below.

Jan 1, 1970

By Sukumar Bandopadhyay

The Marine Mining Technology Center (MMTC) was authorized by an act of the United States Congress in 1996. The act established several marine minerals research centers in the U.S., including one at the University of Alaska Fairbanks. The Alaska MMTC is funded through the Minerals Management Service of the Department of the Interior and concentrates its research in the Artic seas region. Recent research projects at the MMTC include, among others, a geographic information system (GIS) application to the gold placer deposits offshore Nome, Alaska, a marine gold placer reserve definition model using a neural network, and development of an expert system model for submarine tailings disposal (STD) planning and decision-making. Although geostatistics remains the most prevalent method used today for ore grade estimation, researchers have made numerous improvements over the years for accurately predicting grades of ore using various other estimation techniques. As a result, the neural network has recently emerged as an alternative to geostatistics for grade estimation purposes. A neural network is a non-linear, model-free estimator that is well-suited for noisy and/or sparse data environments. Neural networks perform best with non-linear spatial data, contrary to the stationary assumption of ordinary Kriging techniques. In the marine placer ore reserve definition model, the neural network technique was used to estimate placer gold reserves offshore Nome. This study demonstrated the utility of neural network modeling over other traditional ore reserve estimation methods, especially for high cut-off grades. Some of the relevant issues of neural network modeling were also addressed in the study. In testing, the neural network produced results that were close to those from geostatistical techniques including Kriging, polygon, and inverse distance methods.

Jan 1, 2004

Allan Gibb, Queensland, Australia (communication to the Secretary*):—It is gratifying that Mr. Edward Keller,&apos; who has clone so much work elucidating the principles of copper-metnllurgy, should have subjected only that portion of our investigation relating to magnetic iron oxide to anything like adverse criticism. Situated as we were on a rnining-camp, it was impossible to undertake work that would cover all the intricate possibilities connected with the materials commonly classed as matte. We therefore began the investigation on matte produced from ores that contained no other components than the sulphides of copper and iron that would be likely to enter into the composition of matte. We were, accordingly, free from ally possibility of magnetic iron oxide entering the matte, unless it were formed in the operation itself. The work done on this subject was incomplete, and the results thereof recorded as giving only negative evidence as to the absence of magnetic iron oxide. It would have been better to have left the matter open, as " not proven." The trend of the investigation was to account for the insufficiency of sulphur in matte to satisfy the requirements of ferrous and cuprous sulphides. So far as our deductions are correct, it was unnecessary to consider the (in our samples) somewhat doubtful constituent, magnetic iron oxide. The mattes that mere investigated were among those mentioned by Mr. Keller, " which scarcely show ally magnetic property." The samples showed no such property. There was no unaccountable variation in the specific gravity, which only varied, as would be expected, with the proportion of ferrous sulphide. I have had experience in smelting ores of which the gangue was a siliceous magnetite, both in blast- and rever-beratory furuace. The mattes so produced contained magnetic

Jan 1, 1908

By M. R. Annis, P. H. Monaghan

The control of mud properties affords two practical means of tnitigating pipe sticking caused by differential pressure: (I) teducing weight and, therefore, differential pressure; and (2) reducing the friction berween the pipe and mud cake. This paper describes investigation of the second of these—the friction between the pipe and the mud cake. Friction between a steel plate and a mud cake, held in contact by a differential pressure, was measured in the laboratory while maintaining a constant area of contact. Experiments were performed to determine how this friction varied with changes in mud composition and with changes in experimental conditions such as the differential pressure, time of contact of plate and mud cake, and filter-cake thickness. It was found that the apparent coefficient of friction, or the "sticking" coeficient, was not a constant; instead, it increased with increased time of contact between plate and mud cake, and with increased barite content of the Mud. The sticking coeficient varied from about 0.05 to 0.2 afer 20 , and eventually reached values of 0.1 to 0.3 after two Hours. Quehracho or ferrochrome lignosulfonate reduced the sticking coefficient at short .set times but did not reduce the maximum value. Carboxy-~t~etlz~lcellulose had no effect on the sticking coeficient. Emulsification of oil in the mud reduced the sticking coefficient. Some oils reduced the sticking coefficient to about one-third of its Value in the oil- free base mud, while other oils reduced it only slightly. Addition of certain surfactants with the oils further reduced the sticking coefficient. Spotting a clean fluid over the stuck plate caused a reduction in sticking coefficient only if the differential presslrrr was reduced, either temporarily or- permanently. INTRODUCTION Often during drilling operations the drill string becomes stuck and cannot be raised, lowered, or rotated. This condition can be brought about by a number of causes, such as sloughing of the hole wall, settling of large particles carried by the mud, accumulation of mud filter cake during long stoppage of circulation and, finally, sticking by pressure of the mud column holding the pipe against the filter cake on the hole wall. This paper is concerned with the last-mentioned phenomenon. Helmick 2nd Longley&apos; in 1957 suggested that a pressure differential from the wellbore to a permeable formation covered with mud cake could hold the drill pipe against the borehole wall with great force. This situation occurs when a portion of the drill string rests against the wall of the borehole, imbedding itself in the filter cake. The area of the drill pipe in contact with filter cake is then sealed from the full hydrostatic pressure of the mud column. The pressure difference between the mud-column pressure and the formation pressure acts on the area of drill pipe in contact with the filter cake to hold the drill pipe against the wall of the borehole. Helmick and Longley also presented laboratory cxperiments which showed that the force required to move steel across a mud cake increased with increasing differential pressure and with the time the stcel and mud cake had been In cuntact. Their data indicated that replacing the bulk mud with oil reduced the force required for movement. Field evidence was rcported that spotting oil over the stuck interval sometimes freed the pipe. Outmans- in 1958 presented a theoretical paper which described the sticking mechanism and explained the increase of sticking force with time with equations derived from consolidation theory. Since publication of these papers, there has been interest in the differential pressure sticking of drill strings, and several mud additives to reduce sticking or special equipment to free stuck pipe have been proposed."" Haden and Welch" have recently reported laboratory evidence showing that the composition of the filter cake influences the force necessary to move steel on the filter cake. There seems no doubt that differential pressure sticking is a real phenomenon and that its severity depends on the magnitude of the pressure differential across the mud cake, the area of contact and the friction between pipe and mud cake. The mud weight required to control a well is determined by the highest formation pressure in the well: hence, the magnitude of the differential pressure opposite normal or subnormal pressure formations cannot bc reduced. The area of contact may be minimized in several ways (control of filter-cake thickness, use of stabilizers and spirally grooved drill collars), but there arc practical limitations which prevent reduction of contact area from becoming a complete solution of the problem. However. the mud composition might bc altered to reduce the friction between pipe and mud cake. This paper presents quantitative measurements of the friction between steel and mud filter cake and shows how the friction varies with mud composition for given experimental conditions.

By H. A. Bourne, E. L. Haden, D. R. Reichmuth

A circumferential-toothed bit with novel tooth form gave improved penetration performance. In this design the exterior flank of all teeth were vertical when in rolling contact with the hole bottom. Rock chips were generated by the interior flank of the tooth displacing the rock inwardly and downslope toward the center of the hole. A unique two-cone laboratory bit assembly enabled evaluation of numerous cone and tooth configurations. Some of the variables investigated, in addition to weight on bit, rotary speed and rock type, were tooth interference, percent tooth, hole bottom angle, attack angle and relief angle. Most tests were conducted dry on a brittle synthetic sandsone or a ductile quarried limestone. Tooth configurations were found to be more significant in the ductile material. This was attributed to the deeper tooth penetration before rock failure. These studies showed that the attack angle (angle beween interior flank of the tooth and rock surface) was the controlling variable; changing the tooth configuration from the assymetric or semi-wedge to the more conventional symmetric or wedge form reduced penetration performance; and penetration performance of circumferential-type cutters was directly proportional to rotary speeds up to 200 rpm. INTRODUCTION Much of the published literature on rock-chisel interactions describe experiments wherein symmetrical wedges are vertically loaded or impacted against a smooth rock surface.1-6 are is usually taken to insure that the indentation is not made near the edge of the rock specimen less erroneous data be obtained. The literature describes relatively few studies in which the investigator deliberately attempted to take advantage of an edge or free surface. In contrast, anyone who chips ice or breaks up a concrete sidewalk almost always works near an edge. Chisel "indexing," which has been considered by some investigator1,2,6,7 makes limited application of an edge or free surface. Probably the best documented investigation into applying this idea to drilling was that of Drilling Research Inc. at Battelle Memorial Institute.&apos; Their "annular wing" percussion bit consisted of paired asymmetric chisels oriented so as to produce and move chips to the center of the hole. They predicted that the lowest energy requirement for chip generation would be achieved with a stepped hole bottom having a median angle of 45" to the horizontal. Results from limited tests showed that approximately 50 percent of the rock fragments were large and semi-circular in shape, as would be expected by a chisel impact near an edge. The remaining 50 percent were fine chips produced by the chisels in re-establishing the steps or ledges. Initial penetration rates with this bit were high, but they rapidly decreased. This was the result of excessive tooth wear caused by the constant friction on the gauge surfaces. The basic idea — circumferentially placed asymmetric chisels — still appears to have merit. If the concept could be applied to a rolling cutter bit, two of the shortcomings of the fixed chisel design could be overcome: (1) reduction in tooth friction, and (2) greatly increased cutter surface. Adapting asymmetric chisels to cutters rolling on an inclined hole bottom is restricted by bit geometry. The basic elements of roller rock-bit construction prevents the practical attainment of a 45" hole bottom angle. Nonetheless, experimentally it was considered desirable to investigate the influence of hole bottom angle to at least 40". This paper describes the laboratory studies conducted in evaluating the circumferential-toothed roller cutter rock bit. EXPERIMENTAL APPARATUS AND PROCEDURE BIT ASSEMBLY The cost of constructing a sufficient number of conventional three-cone rock bits to investigate circumferential cutter performance was prohibitive. Instead, a novel two-cone laboratory assembly which used an external bearing system was designed and constructed. The external bearings made it possible to alter the journal bearing angles and thus allow a wide flexibility in cutter configuration. Fig. 1 shows the laboratory bit assembly, the various bearing mount plates and the appropriate roller cutters for drilling shallow holes having hole bottom angles of 0, 10, 20, 30 or 40". The bit was limited to a drilling depth of 1 1/2 in. at the gauge teeth and a hole diameter of 43/4 in. This more or less intermediate size bit was chosen because it gave a more realistic match between bit teeth and the rock than would a microbit. Also, the rock sample size required was convenient and easy to obtain. CIRCUMFERENTIAL CUTTERS The tooth configuration used in our initial studies is shown in the upper half of Fig. 2. All cutters used in this series had the same tooth form — 43" included tooth angle, 2" positive relief angle and a horizontal tooth flat width of 1/32 in. Each cone cuts alternate rows except for the gauge row. The row-to-row spacing in view was 1/4 in. Static loading tests conducted earlier with asymmetrical chisels had been used to establish this spacing. These tests showed energy requirements for chip production increasing rapidly as the distances to the edge increased beyond

By Hugh H. Waesche

Of all the military services, the Signal Corps is the most concerned with piezoelectric minerals because of its function as a supply service to the strategic and tactical military forces. Consequently this paper is written from the point of view of one associated with that organization. The Signal Corps is responsible for the research, development, and supply of communications, radar, and components to the using services of the Department of the Army and to some extent the Other branches of the National Defense Department. Their work therefore includes the study of the sources* characteristics, and application of quartz and other piezoelectric materials. These materials have become a vital consideration in strategic planning and are essential for efficient tactical operation by all the Armed Forces. The Signal Corps at the beginning of world War 11 Was respon-sible for both Army Ground and Air Force electronic equipment. Since that time this Army service organization has probably done more in the development of frequency control devices using piezoelectric materials than any other group. The U.S. Department of the Interior, Bureau of Mines, Minerals yearbook of 1945, shows that during the four war years, 1942 through 1945, 9,598,-410 Ib of quartz crystal were imported for all uses and of this total, 5,168,000 lb were consumed to produce 78,320,-000 crystal units for electronic application. Other government records confirm these data which conclusively show that approximately 53 pct of the crystalline quartz imported was consumed in the production of electronically applied quartz crystal units. It may be assumed that some effort was made to maintain a stockpile over demands for all purposes. and this would mean that the actual percentage of quartz used electronically was considerably over the 53 pct figure. These data only emphasize that electronic application of crystalline quartz was the greatest requirement, and per- haps the actual value in this application to national defense is many times greater in importance than is apparent on first inspection. Current electronic research and development programs of the Armed Forces are planned around the fundamental use of piezoelectric minerals for frequency control and this at present, at least, means quartz. Definition and Early Development The word piezoelectricity is formed from combination of the Greek word "piezein". meaning "to press," and "electricity." It is that property shown by numerous crystalline substances whereby electrical charges of equal and opposite value are produced on certain surfaces when the crystal is subjected to mechanical stress. It appears to be intimately associated with the better known property, pyro-electricity and in fact, the two may be manifestations of the same phenomeuon. This property was discovered by Pierre and Jacques Curie in quartz, tourmaline, and other minerals in 1880 while studying the symmetry of crystals. The converse effect, that is, mechanical strain in the crystal when placed in an electrical field, was predicted by the French physicist, G. Lippman, in 1881, and verified by the Curies almost immediately. As has been the case with many discoveries of similar character in the basic sciences, not much attention was paid to this property for man)- years except as an entertaining curiosity. Between 1890 and 1892 a series of papers was published by W. voigt in which the theoretical physical properties were put into mathematical form. The first practical application of piezoelectricity occurred during World War I when professor P. Langevin of France used quartz mosaics to produce underwater sound waves. The same mosaics were used to pick up the sound reflections from submerged objects which were in turn, amplified by electronic means and used to determine the distances to such objects. This device was intended for use as a submarine detector but development was not completed in time for war service although it was used later for determining ocean depths. About the same time, A. M. Nicholson, of Bell Telephone Laboratories, developed microphones and phonograph pickups using Rochelle salt crystals. A major step in the application of piezoelectric quartz came in 1921, when professor W. G. Cady, of wesleyan university, showed that a radio oscillator could be controlled by a quartz crystal; from that date, this use of quartz has increased steadily, reaching its peak in world war 11 as is shown by the figures previously given. Essentially all American electronic equipment for communication, navigation, and radar, utilized quartz crystals for the exacting frequency control required. Crystalline Minerals with piezoelectric Properties QUARTZ Hundreds of piezoelectric crystalline materials are known, most of which are water soluble. Of these, quartz appears to be without a peer for electronic frequency control. Unfortunately, the quartz must be of superior quality. It must be free of mechanical flaws, essen-tially optically clear, free of both Brazil and Dauphiné twinning and must be, for average uses, over 100 g in weight. Because of these stringent requirements, raw quartz of the quality desired is of rare occurrence. In addition to quartz, several other naturally occurring crystalline materials are known to have the piezoelectric property and could perhaps be substituted for quartz in some applications. These

Jan 1, 1950

By J Price, M. R. Geer, H. F. Yancey

COAL mine operators recognize coal as an abrasive material, because the wear of drilling, cutting, and conveying equipment is reflected as a cost item for replacement of parts. Similarly, industrial consumers of coal experience abrasive wear on all coal-handling equipment. Operators of pulverized fuel plants are doubtless most keenly aware of the abrasiveness of coal, because under the high contact pressures developed between coal and metal in pulverizers, abrasive wear is increased many fold. Moreover, experience in operating pulverized fuel plants has demonstrated that some coals are much more abrasive than others. Hardgrove&apos; stated that maintenance costs entailed by the wear of grinding elements is often a more important variable than the cost of the power required to pulverize different coals. Craig2 also reports that one coal may cause pulverizer parts to wear several times faster than another. It is apparent, therefore, that those concerned with pulverizing coal could profitably employ a method for estimating the abrasiveness of different coals, just as they utilize standard tests for thermal value, grindability, and ash-fusion temperature to assist in selecting the most suitable and economical coal to use in a particular plant. The objective of this investigation was to develop a test procedure that would be suitable for general use in estimating the abrasiveness of coals. However, few, if any, of the standard tests now used for evaluating the properties of coal are the product of a single investigation or the result of a single investigator&apos;s efforts. Rather, in each case, a testing procedure was devised by one investigator, used by others on a wider variety of coals, and finally refined completely as the result of the joint efforts of a number of interested people. Thus, the test procedure for estimating abrasiveness developed in the course of this work may not be refined sufficiently in its present form for general use, but it may serve as the starting point from which an acceptable test procedure can be developed. The method has been used thus far on only about a dozen coals, and there has been no opportunity to attempt a correlation between experimental results and actual plant experience. Only wider use of the procedure by other investigators and correlation with plant experience can determine to what extent the method will have to be modified to render it suitable for general application. Test Method Although the literature contains no record of an attempt to devise a method for estimating the abrasiveness of coal that could be used industrially, several investigators have tested properties of coal that are closely related to its abrasiveness. The abrasiveness of a material generally is considered to be related to its hardness, and hardness tests for coal have been employed by Heywood,&apos; O&apos;Neill," and Mathes. Also, the resistance of coal to abrasion, a property that presumably is related to the abrasiveness of coal, was measured by Heywooda and by Simek, Pulkrabek, and Coufalik.2 11 these investigators tested only individual pieces of coal. Since coal is a heterogeneous material having components of varying properties, tests of this type can yield results having little more than academic interest. Only a test method that utilizes a representative sample of coal can give results that are useful industrially. The abrasion tests used for various other materials have been considered for adaptation to testing the abrasiveness of coal. The tests used for metals,7-9 paving and flooring,&apos;" and rubber," cannot be used because coal is not sufficiently abrasive.~ The present experimental work was begun before World War II and was conducted by three research fellows"&apos;" working under a joint agreement between the University of Washington and the Bureau of Mines. After a great deal of preliminary work with a variety of apparatus and materials, a test procedure was developed which consisted of rotating a test disk 2Yz in. diam in a steel mortar containing the coal sample. The shaft carrying the test disk at the lower end and a 100-lb load on the upper end was free to move vertically. The bed of coal in the mortar was kept fluid by low-pressure air admitted through a port near the bottom of the mortar. Measurable wear on an Armco iron disk could be obtained in this test procedure, but, despite extensive efforts to eliminate them, several major disadvantages remained in this test method. First, with most coals the amount of wear on the iron disk did not exceed a few milligrams. Second, a single type of disk was not applicable for all coals. A smooth iron disk gave satisfactory results with both bituminous and sub-bituminous coals, but hardly any wear with anthracite or coke. A disk having studs or projections gave more satisfactory abrasion losses with anthracite and coke and presented no operating difficulties with free-burning bituminous and sub-bituminous coals. It could not, however, be used with caking coals because these coals formed a

Jan 1, 1952

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country&apos;s sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch&apos;s hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world&apos;s sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO2 oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO2 can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1 ½ ¢ per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past year that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 ½ million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country&apos;s sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch&apos;s hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world&apos;s sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ¾ ton of high-grade pyrite and results in ½ ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15¢ per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1952

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country&apos;s sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch&apos;s hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world&apos;s sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By H. J. Tschirner, L. P. Davidson, R. K. Carpenter

HE electrolytic zinc plant of the American Zinc Co. of Illinois, at Monsanto, was expanded in 1943 to a capacity of 100 tons of slab zinc daily. This capacity was not attained because of inability of the leaching plant to deliver an adequate amount of solution for electrolysis. Changing the leaching method so that the acid was added to the roasted zinc material reversed the usual procedure and made it possible to attain the desired capacity. The conditions which prevented satisfactory work before this change and the difficulties which arose in reversing the usual leaching procedure are described. The "reverse" leach operation as now practiced is carried out as follows: All the calcine to be leached is fed continuously to a slurry mixing tank. About one third of the acid to be used is fed to the tank with the calcine. The slurry is discharged continuously to a Dorr duplex classifier in closed circuit with a Hardinge mill. The classifier overflow is pumped to any of six leaching tanks where the leach is completed. A finished leach is discharged through Allen-Sherman-Hoff pumps to Dorr thickeners, from which the overflow goes to the zinc dust purification and the underflow to vacuum filters. This change in leaching procedure from the usual one of adding calcine to a large amount of acid made it possible to provide an adequate amount of purified solution to the electrolyzing division and at the same time filter and dry all the residue produced. Operating savings in reagents and better metallurgical recoveries were also important benefits. The original flowsheet of the leaching plant provided leaching, sedimentation of the insoluble residue, and purification of the neutral zinc sulphate solution with zinc dust. The thickened residue was filtered and washed. The purification cake of excess zinc dust, precipitated copper and cadmium, and any insoluble residue was filtered off on plate-and-frame duplex classifier. Settlement in the thickeners was inadequate and the suspended solids in the thickener overflow gave rise to filtration difficulties after the zinc dust purification. Further, the filtration and washing of the leach residue was poor, and it became necessary to pump a large amount of unwashed or poorly washed residue to storage ponds outside the plant building. Two causes of the poor settling and filtration were determined: Soluble silica and ferrous iron in the calcine treated. The latter was a result of poor roasting and with more experience ceased to be a major problem. The silica was a normal constituent of the feed and the working out of the problem became a matter of controlling its solubility. The obvious method to render the silica insoluble was by intensive roasting. This, however, met with total failure as such roasting resulted in silicates, probably zinc, soluble in the 13 pct acid used for leaching. Attempts were made to coagulate the fine gelatinous slime with addition agents. Glue, lime, starch, beef-blood serum and others were tried without success. All the suggested tried-and-tested means of operating the thickeners gave no consistently good results. Variations in leaching time, in addition of calcine to the leaching tanks, "conditioning" of the pulp by prolonged agitation, immediate discharge of the leach upon completion to avoid breaking up flocs were all tried and given up as ineffective. Byron Marquis, of Singmaster and Breyer, worked with the plant staff on a beaker scale until a leaching procedure was developed which gave consistent results and a promise of overcoming the difficulties which had plagued the plant operation. It was suggested that the difference in solubility of silicates and zinc oxide in sulphuric acid could be made use of in a leaching method where the acidity was controlled carefully. Such control is possible when acid is added to a slurry of calcine. This process reverses the normal procedure of adding calcine to a vessel of acid, hence the term "reverse leach" was applied. In this way, the overall acid concentration can be kept very low. In the tests made, it did not exceed 0.05 g per liter free sulphuric acid. Numerous advantages were realized when no silicates were taken into solution and later precipitated as a bulky gel. The gel had made reasonable thickening and filtration of the leach pulp and practical drying of the residue impossible. When the gel was eliminated, thickening rates were increased as much as five times. The volume of residue after thickening represented about 10 pct of the total leach pulp and had been as high as 95 pct when the gel was present. The thickened pulp was filterable and the filtered cake was dried readily after washing. The zinc extraction from the calcine was slightly lower. This was more than compensated for by the increase in zinc recovered in solution from zinc which had been trapped in the gelatinous residue. The amount of copper recovered was lower. However, the amounts of other impurities, such as arsenic, antimony, and germanium, taken into solution were lower. This was particularly true of antimony. Since the inception of reverse leaching, no concentrates have failed to yield solutions free of antimony even when present in the calcine to the extent of 0.2 to 0.3 pct. Oxidation of ferrous iron is a problem of reverse leaching. Ferrous hydrate does not precipitate at pH 5.3 to 5.4 where a leach is finished. The usual oxida-

Jan 1, 1952

By R. J. Fruehan, F. D. Richardson

Electrochemical measurements have been made of the activity of oxygen in copper and its alloys with silver and tin at 1100" and 1200°C. The galvanic cell used was Pt, Ni + NiO/solid ellectrolyte/[O] in metal, cermet, Pt The results do not support any of the equations so far designed for predicting the activities of dilute solutes in ternary solutions from activities in the corresponding binaries. If, however, a quasichemical equation is used with the coordination number set to unity, agreement between observed and calculated activities shows that this empirical relationship can be useful over a fair range of conditions. SEVERAL solution models have been proposed for predicting the activity coefficients of dilute solutes in ternary alloys from a knowledge of the three binary systems involved. Alcock and Richardson1 have shown that a regular model, and a quasichemical model,&apos; in which the dissolved oxygen is coordinated with eight or so metal atoms, can reasonably predict the behavior of both metal and nonmetal solutes when the heats of solution of the solute in the separate solvent metals are similar. But when this is not so, neither model gives useful predictions unless coordination numbers of one or two are assumed. Wada and Saito3 subsequently adopted a similar model to derive the interaction energies for two dilute solutes in a solvent metal. Belton and Tankins4 Rave proposed both regular and quasichemical type models in which the oxygen is bound into molecular species, such as NiO and CuO in mixtures of Cu + Ni + 0. However, their models have only been tested on systems in which the excess free energies of solution of the solute in the two separate metals differ by a few kilocalories. Ope of the reasons why more advanced models have not been proposed is because few complete sets of data exist for ternary systems in which the solute behaves very differently in the two separate metals. For this reason measurements have been made of the activities of oxygen dissolved in Cu + Ag and Cu + Sn. Measurements on both systems were made by means of the electrochemical cell, Pt, Ni + NiO/solid electrolyte/O(in alloy), cermet,Pt [1] The activity of oxygen was calculated from the electromotive force and the standard free energy of formation of NiO, which is accurately known.5 Before investigating the alloys, studies were made of oxygen in copper to test the reliability of the cell and to check the thermodynamics of the system. Of the previous studies those by Sano and Sakao,6 Tom-linson and Young,7 and Tankins et al.8,7 have been made with gas-metal equilibrium techniques; those by Diaz and Richardson,9 Osterwald,10 wilder," Plusch-kell and Engell,12 Rickert and wagner,13 and Fischer and Ackermann14 have been made by electrochemical methods. EXPERIMENTAL The apparatus employed was the same as described previously,9 apart from slight modification. The molten sample of approximately 40 g was held in a high grade alumina crucible 1.2 in. OD and 1.6 in. long. The solid electrolytes were ZrO2 + 7½ wt pct CaO and ZrO2 + 15 wt pct CaO; the tubes 4 in. OD and 6 in. long were supplied by the Zirconia Corp. of America. They were closed (flat) at one end. In one experiment with Cu + O, both electrolytes were used in the cell at the same time. The reference electrodes inside the electrolyte tubes consisted of a mixture of Ni + NiO. They were made by mixing the powdered materials and pressing them manually into the ends of the tubes, with a platinum lead embedded in them. The tubes were then sintered overnight in the electromotive force apparatus at 1100°C. By sintering the powders inside the tubes (instead of using a presintered pellet9) better contacts were obtained between the electrolyte, the powder, and the platinum lead. Troubles arising from polarization9 were thus much reduced. The electromotive force was measured by a Vibron Electrometer with an input impedence of 1017 ohm; the temperature was measured with a Pt:13 pct Rh + Pt thermocouple protected by an alumina sheath. The couple was calibrated against the melting point of copper. The cermet conducting lead of Cr + 28 pct Al2O3, previously found to be satisfactory9 for use with Cu + 0 was also found satisfactory with Cu + Ag + 0 and Cu + Sn + 0. Superficial oxidation was observed, but it did not interfere with the working of the cell. The reaction tube containing the cell was closed at each end with cooled brass heads and suspended in a platinum resistance furnace. The tube was electrically shielded by a Kanthal A-1 ribbon which was wound round it, and the ribbon was protected by a N2 atmosphere between the furnace and the reaction tube. The cell was protected by a stream of high purity argon which was dried and passed through copper gauze at 450°C and titanium chips at 900°C. All the metals used were of spectrographic standard. Procedure. In studies of the system Cu + 0, be-

Jan 1, 1970

By H. F. Ramstad, F. D. Richardson

The equilibria (a) SiOz +Hz =SiO +H20 and (b) Si + SiO, = 2Si0 have beet1 studied at temperatures of 1425"to 1600°C ad 1310°to 1485°C respectively. The stattdard free energy changes for the tzrro reactions are given by the equatiotts Combination of the results for both equilibria leads to tiotz removes certain anomalies in existing high-terlzperature data for equilibria involving silica and silicon in iron. In many metallurgical processes and in many laboratory investigations silicon monoxide undoubtedly plays an important role. It is unfortunate therefore that wide differences exist between the results obtained by different investigators1-7 in their studies of such equilibria as In an attempt to put our knowledge of SiO on a surer basis, an exhaustive study has been made of equilibria [I] and [2] at temperatures ranging from 1300" to 1600°C. Reaction [I] was studied by measuring the amounts of silica which could be condensed from streams of Hz or Hz + HzO which had previously been brought into equilibrium with silica at temperatures ranging from 1425" to 1600°C. Reaction [2] was studied by measuring the material that could be condensed from streams of Hz or argon which had been brought into equilibrium with mixtures of silicon and silica at temperatures ranging from 1310" to 1485°C. EXPERIMENTAL Materials. The silicon was "superpure" grade and contained less than 0.1 pct impurities. The silica was prepared from pure mineral quartz; this was crushed and treated with concentrated hydrochloric acid to remove particles of iron, washed with water, and finally dried at 120°C. For the hydrogen + silica reaction, the silica was sized to —20+100 mesh. For the silicon + silica reaction, the two materials were ground to a fine powder in an agate mortar. The hydrogen and argon were commercial oxygen-free gases. The gas streams were controlled with capillary flow meters and the volumes were measured by wet gas meters. After passing through the meters, the gases were partially dried by silica gel. The hydrogen for the HZ + SiOz reaction was then passed through palladised asbestos at 300°C and dried with magnesium perchlorate. The efficiency of oxygen removal was checked throughout the experiments by passing the gas over an electrically heated strip of nichrome, used as an indicator as described by Rathman and de itt.&apos; When mixtures of HZ + Hz0 were required, the partial pressures of water vapor (1.8 to 22 mm) were obtained by passing the hydrogen through oxalic acid dihydrateg&apos; lo held at various controlled temperatures, O.l°C, by means of a water bath. The hydrogen for the Si + Si02 reaction was purified by passing it over a mixture of 3 parts of magnesium to 5 of lime heated to 600°."l1 u The argon for this reaction was passed through titanium powder (-3/16 in. + 100 mesh) heated to 900°C. The nitrogen used to prevent the reaction products escaping from the condenser (see later), was deoxidized by copper or iron at 600°C. All these gases were finally dried with magnesium perchlorate. Furnace, Temperature Contr01, and Measurement A molybdenum resistance furnace was used for both sets of experiments. The reactions were conducted inside a high-grade alumina tube, 36 in. long and 1 in. in diam as indicated in Fig. 1. With this arrangement an even temperature zone (2"C) 4 cm long was satisfactorily obtained. The temperatures were kept constant by means of a proportional controller actuated by a Pt-Pt 13 pct Rh thermocouple. This was placed between the two alumina tubes, so that the temperature at the junction was 1400" to 1450°C. Up to 1485"C, the temperatures were measured with Pt-Pt 13 pct Rh thermocouples. For higher temperatures an optical pyrometer was used, this being sighted (through the glass window 1 in Fig. 1) on the end of the alumina tube, that held the SiOz or Si +SiOz mixture, 10 in Fig. 1. The optical pyrometer was recalibrated whenever a change was made in any part of the apparatus situated in the hot zone. Successive readings with the optical pyrometer were reproducible to within 1"C. Equilibrium Apparatus and Procedure. Hydvogen and Silica Reaction. The apparatus is shown in Fig.

Jan 1, 1962

By Charles R. Stewart, William W. Owens

Reservoir performance predictions based on laboratory core test data assume that fluid flow is laminar for the laboratory test. A study has been made to determine the validity of this assumption for laboratory tests on various types of porosity found in producing limestone formation. Data are presented which show that turbulence and slippage can occur during laboratory tests on hetero-geneous-type porosity limestones, thus causing serious errors in measured single-phase permeabilities and two-phase relative permeability characteristics. In single-phase flow tests it is possible to eliminate turbulence and correct for slippage or to eliminate both factors by controlling test conditions. It is not always possible to control test conditions and thereby eliminate turbulence and slippage in two-phase .flow tests. A correction method is presented which can be used to calculate the true two-phase laminar flow relative permeability characteristic even though furbulence and slippage exist. .INTRODUCTION It is customary to make use of Darcy&apos;s law and modifications of this law, together with laboratory data on formation core samples to predict the performance of producing reservoirs. Such predictions are based on an assumption that fluid flow is in the laminar or streamline region for the laboratory test. It was the purpose of this inves- tigation to determine the extent to which turbulent flow may occur in laboratory fluid flow tests on hetero-geneous porosity limestones. Considering that turbulent flow conditions might exist in some laboratory fluid flow tests, additional emphasis was placed on the development of a method to correct for turbulence when laminar flow conditions could not be attained. FLUID FLOW CONCEPTS FOR POROUS MEDIA The Influence of Pore Geometry on Fluid Flow One of the more important factors influencing fluid flow in porous media is the geometry of the pore space which includes such characteristics of the pores as size, shape, distribution, roughness, uniformity, etc. In general, oil- and gas-producing formations can be divided into two broad types on the basis of pore geometry. One has been called sandstone-type porosity media, which is characterized by a small range in pore size, uniformity in shape of the pores, smooth pore surfaces and a regular and uniform distribution of pores. The other type has been called heterogeneous porosity media and is usually limited to the dolomites and limestones. This type is characterized by a wide variation in the size, shape, and distribution of the pores and rough, irregular pore surfaces. It is therefore apparent that conditions are much more favorable for turbulent flow* in heterogeneous-type porosity media than in sandstone-type porosity media. Interrelationship Between Turbulence and Gas Slippage In studying the problem of turbulent flow in laboratory tests on porous media, it is necessary to be aware of the interrelationship between slippage and turbulence for gas flow. As a result of slippage or the Klinken-berg effecta, apparent perrneabilities to gas are greater than the true value because there is no stationary layer of gas in contact with the walls of the flow channels. Gas slippage decreases as the mean free path of the gas molecules decreases. Since the mean free path of any gas decreases with increasing density, increases in static pressure result in lower apparent gas permeabilities. However, a reduction in gas permeability can also be due to turbulence. Therefore, in studying only turbulent flow in porous media, it is necessary to hold gas density, and slippage, constant or to reduce slippage to a negligible value by operating at high static pressures. Presentation of Laminar and Turbulent Flow Data A graphical relationship between permeability and a pseudo-Reynolds number, N,, will be used to show the two types of fluid flow, i.e., laminar and turbulent. The usual graphical method for such a description has been the use of friction factor-Reynolds number charts4. On such a logarithmic diagram, the laminar region appears as a straight line having a slope of 45 degrees. As the friction factor decreases and the Reynolds number increases, the turbulent region is reached and appears as a deviation from the 45-degree slope line. However, in petroleum engineering literature resistance of por-

By Mikhail D. Ageev

The paper concerns with some questions of Autonomous Underwater Vehicle (AUV) control system design, which allows a motion at rough undersea terrain. The AUV advantages reveal itself completely during exploration of such objects as seamounts and reefs area. These objects are characterized by steep slopes (30 degrees and above) and chaotic situated obstacles (rock debris, pinnacles, escarpments). Uses of other technical facilities for operating in such formations are difficult, dangerous or impossible. These regions take an interest due to the presence of iron-manganese crusts. Modern AUV can estimate density of mineral by means of still camera or video recorder. For this reason the vehicle has to move with the distance of 3-5 meters from seabed. Under such conditions micro- and mesorelief presents a considerable danger for AUV. The AUV can bypass underwater obstacles in the various ways. However, in any case AUV motion will be characterized by intensive maneuvering on various speed (from 0 up to 1.5 m/s), circular change of tractive force vector and, hence, a circular flow about the hull. So, AUV is equipped with the propulsion system, which provides free modes of motion. AUV control system (CS) has tri-level structure to allow a motion in a rugged seabed terrain. According to the generally accepted categorizations, CS includes execution, coordination and strategic levels. The strategic level contains the goals of AUV mission. Obstacles avoiding and route choice are realized by means of simultaneous maneuver in vertical and horizontal planes. Sonar system is used for data acquisition about obstacles. It consists of a few echo-sounders with fixed directivity pattern. The route choice algorithm operates at two lower levels of CS. Basic (?reflex?) functions of spatial motion are realized within the execution level and considered in the paper. The choice of motion modes is carried out at the coordination level of system. All accessible navigating information (the data from flight sensors and from multichannel echo-sounder systems) is used for this purpose. In addition the bypass of characteristic types of obstacles is carried out with the help of basic motion modes combination. The modeling results of bypass of various typical obstacles, and also their combinations are described in the paper.

Jan 1, 2011