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George IV. Goetz, Milwaukee, Wis.: Mr. Campbell deserves much credit for his interesting paper. The literature of the development of the open-hearth process is distributed in many technical journals, and is difficult of access to many engineers. Since the development of the basic process, the open-hearth furnace has been raised from a subsidiary to an independent position. Certain high grades of soft steel can be produced in the basic open-hearth furnace, which cannot be produced by any other known process. In the United States the charges for the acid process generally consist of either scrap and pig-iron or so-called wash-metal and scrap : the pig-and-ore process, as carried out in England, is practiced but very little. The introduction of the basic process has made it possible to use about four times as much pig-iron with scrap as can be used in the acid mixtures. The availability of a cheap pig-iron, very low in silicon and sulphur, and carrying several per cent. of manganese, is one of the reasons why the basic open-hearth process has developed so rapidly in Germany, as regards the output per furnace and the excellent quality of the material produced. The acid Bessemer steel has an inherent stiffness, which is plainly shown when spring steel is made by the Bessemer and by the openhearth processes. In order to get them of about equal stiffness, the open-hearth steel must have about 0.60 per cent. carbon ; whereas, the Bessemer steel must then have ahout 0.45 carbon, the remaining constituents being apparently about the same in each product. It has proved impracticable, with American coke-pig-iron, to interrupt blowing at a certain point in order to make hard steel; and experience has therefore shown that it is best to eliminate nearly all the carbon, and then to recarburize to the desired point. This practice

Jan 1, 1894

By H. Kahata, I. Iwasaki, Y. Takahasi

A combined process of segregation roasting-flota-tion or magnetic separation was developed to recover nickel from lateritic iron ores and oxidized nickel ores. An ore was mixed with a halide, such as sodium or calcium chloride, and a solid reductant, and roasted, whereby the nickel content was segregated out as discrete metallic particles. Either by magnetic separation, or by flotation using a sulfhydryl collector together with a heavy metal salt as an activator, the metallic nickel could be concentrated, Effects of such variables as the types and amounts of the halide and solid reductant addition and of the roasting conditions were extensively investigated. The concentration results on several types of ores from Japan, the Philippines, and New Caledonia are presented. Laterite is a residual product of rock decay in tropical and subtropical regions, and consists mainly of hydrated oxides of iron and alumina. Those derived from serpentine sometimes become enriched in nickel, cobalt, and chromium contents. Oxidized nickel ores are of the similar origin, but their nickel as well as silica contents are higher due to gamierite-type silicates. Apparently, nickel enrichment is more pronounced in the transition region between the surface layer of iron laterite and the underlying serpentine. Iron laterites often contain over 50% iron, and their reserve is reported to exceed 22 billion tons, corresponding approximately to 20% of the world iron ore reserve.1 It is readily apparent, therefore, why the iron laterites have attracted much attention of the steel industry as a potential source of iron ores. Numerous attempts to remove nickel and chromium as byproducts, have been reported in literature.' Worldwide discoveries of high-grade iron ore deposits in recent years have inevitably lessened the interests in iron laterites. However, in view of the tremendous potential for their gross value and supplies of nickel, cobalt, chromium, and alumina along with iron, the problem remains as a challenge. Much of the chromium in laterites is present in the form of relatively coarse particles of chromite, and various classification methods have successfully been applied for its removal.2,3 The nickel, however, is intimately associated with limonite, and ordinary concentration techniques are totally ineffective. Much attention has, therefore, been given to hydro-metallurgical or pyrometallurgical means. Yet, all the known processes appear to have limitations either economically or for the use of products as iron raw materials due to a high residual content of nickel. Some attempts have been made to pre-roast laterites or oxidized nickel ores, particularly with sulfidizing agents, and to remove nickel either by magnetic separation,' or by flotation.4 It is apparent that any such concentration process would require grain growth of nickel during roasting for effective separation. Oxidized copper ores are reported to be concentrated by the segregation roasting-flotation process,5 in which the ores are roasted with a chloridizing agent and a carbonaceous reductant in order to promote the growth of metallic copper through repeated cycles of copper chloride volatilization and its reduction on the solid reductant surfaces. Examination of a vapor pressure-temperature diagram of various metal chlorides indicates that a separation of nickel from iron may be effected by a similar process. In fact, the segregation roasting of laterites was reported to increase the extraction of nickel by the ammonia leach process, as compared to the ordinary reduction roasting.6 In the present article some of the preliminary experimental results of the segregation roasting, followed by either flotation or magnetic separation, on laterites and oxidized nickel ores are reported. The effects of various parameters in the roasting as well as in the concentration steps are discussed. MATERIALS AND METHODS An oxidized nickel ore from Thio, New Caledonia, and three laterites from Homonhon, the Philippines, Wakasa and Miyagawa, Japan, were selected for the

Jan 1, 1967

By E. H. Phillips, A. F. Keenan

6.2-1. Historical Development. Hammering on hand-held drill steel was the earliest type of percussion drilling. It was not until 1838 that Singer developed a steam-operated drilling machine that lifted a drilling bar and let it drop by gravity. J. J. Couch of Philadelphia patented the first percussion drill in 1849. His machine was a steam drill which literally hurled the drill bar at the rock and retracted it by means of a ratcheted wheel. Later J. W. Fowle added automatic rotation of the steel. In 1871, Simon Ingersoll patented a rock drill mounted on a tripod which permitted drilling holes at any angle, vertical to horizontal. Up to the turn of the century, percussion drills in use were piston-type drills. The drill steel was solid and attached to an extension of the drill piston. Hole cleaning was accomplished by the pumping action of the steel in the hole. Down-holes were drilled wet and cleaned with a blow pipe while up-holes were drilled dry, allowing the cuttings to drop out by gravity. The forerunner of our present hammer drill was the free-piston drill which was first patented in 1884, by Sargent. Because solid drill steel was used, it did not successfully clean down holes. J. G. Leyner solved the problem of applying the hammer drill principle to down-hole and horizontal-hole work in 1897. The key was to pass com- pressed air and water through hollow drill steel to clean the hole. Leyner's work on valves and rotation mechanisms increased the blows per minute of hammer drills from 300 or 400 up to 1800 and more blows per minute and made possible the first lightweight drilling machines. The development of automatic feeds followed. 6.2-2. Percussion Drills. TYPES OF DRILLS. There are two types of percussion drills in use today; the piston drill and the hammer drill. In the

Jan 1, 1968

By J. Hunter

LouisiaNa ranks fifth among the oil-producing states. In 1941 the state-wide production was slightly in excess of 118,000,000 bbl. of crude oil and conden-sate—an increase of 15,000,000 bbl. over the total production for the year 1940. During 1941, south Louisiana produced approximately 90,900,000 bbl. of oil, while north Louisiana produced approximately 27,000,000. South Louisiana, therefore, produced about 3¼ times more oil than north Louisiana. During the past year, 574,800,000 M cu. ft. of natural and casinghead gas were produced in Louisiana; thus 1941 gas production exceeded the 1940 production by 104,900,000 M cu. feet. Completion records for the year 1941 show that 1512 wells were completed in or adjoining producing fields. These wells, which were drilled to extend known producing areas or were located within producing areas, are not considered as wildcat wells. Of the total number of field wells drilled, 1152 wells found production; the remaining 360 were junked or dry and abandoned. This means that 76 per cent of the wells drilled were successful. Considering the precarious nature of oil-field development, this percentage of new producing wells is exceedingly satisfactory. Of the total number of field welis com-pleted as producers, approximately 86 wells found new and distinct producing sands. During the past year, 30 new fields were found. Four parishes—namely, Caldwell, Catahoula, Orleans, and Natchitoches— entered the ranks of oil-and-gas-producing parishes in the state during that period. This brings the total number of oil-and-gas-producing parishes in the state up to 48. Thus, 75 per cent of the parishes in Louisiana now contain one or more oil and/or gas fields. Although sufficient data are not available to list the addition reserves found in Louisiana during 1941, it is felt that such reserves undoubtedly exceed twice the total production for that year, or at least 200 million barrels. In reserves, therefore, Louisiana ranks third among the various oil-producing states. South Louisiana Although 16 new fields were found in south Louisiana, the outstanding developments in that part of the state were not limited to new discoveries. In fact, most of the increase in reserve in that area has been most consistently accumulated through extensions of known producing sands and through the finding of new sands in already productive fields, especially at depths below the presently producing sands. Since data concerning most of the new extensions and sands are too extensive for discussion here, the more important developments arc briefly outlined below: Abbeville, Vermilion Parish.—Production extended ½ mile to the west. Anse La Butte, St. Martin Parish.—New oil sand discovered at 9400 ft., thus extending production to east flank. West flank defined by dry holes. Barataria, Jefferson Parish.—New oil sand discovered on east side of field at 7620 ft. New

Jan 1, 1942

By J. Hunter

LouisiaNa ranks fifth among the oil-producing states. In 1941 the state-wide production was slightly in excess of 118,000,000 bbl. of crude oil and conden-sate—an increase of 15,000,000 bbl. over the total production for the year 1940. During 1941, south Louisiana produced approximately 90,900,000 bbl. of oil, while north Louisiana produced approximately 27,000,000. South Louisiana, therefore, produced about 3¼ times more oil than north Louisiana. During the past year, 574,800,000 M cu. ft. of natural and casinghead gas were produced in Louisiana; thus 1941 gas production exceeded the 1940 production by 104,900,000 M cu. feet. Completion records for the year 1941 show that 1512 wells were completed in or adjoining producing fields. These wells, which were drilled to extend known producing areas or were located within producing areas, are not considered as wildcat wells. Of the total number of field wells drilled, 1152 wells found production; the remaining 360 were junked or dry and abandoned. This means that 76 per cent of the wells drilled were successful. Considering the precarious nature of oil-field development, this percentage of new producing wells is exceedingly satisfactory. Of the total number of field welis com-pleted as producers, approximately 86 wells found new and distinct producing sands. During the past year, 30 new fields were found. Four parishes—namely, Caldwell, Catahoula, Orleans, and Natchitoches— entered the ranks of oil-and-gas-producing parishes in the state during that period. This brings the total number of oil-and-gas-producing parishes in the state up to 48. Thus, 75 per cent of the parishes in Louisiana now contain one or more oil and/or gas fields. Although sufficient data are not available to list the addition reserves found in Louisiana during 1941, it is felt that such reserves undoubtedly exceed twice the total production for that year, or at least 200 million barrels. In reserves, therefore, Louisiana ranks third among the various oil-producing states. South Louisiana Although 16 new fields were found in south Louisiana, the outstanding developments in that part of the state were not limited to new discoveries. In fact, most of the increase in reserve in that area has been most consistently accumulated through extensions of known producing sands and through the finding of new sands in already productive fields, especially at depths below the presently producing sands. Since data concerning most of the new extensions and sands are too extensive for discussion here, the more important developments arc briefly outlined below: Abbeville, Vermilion Parish.—Production extended ½ mile to the west. Anse La Butte, St. Martin Parish.—New oil sand discovered at 9400 ft., thus extending production to east flank. West flank defined by dry holes. Barataria, Jefferson Parish.—New oil sand discovered on east side of field at 7620 ft. New

Jan 1, 1942

By D. W. Fuerstenau, D. A. Sullevan

In the experimental work for this comparison of wet and dry grinding, it was found that the size distributions for wet grinding operations are characterized by a constant value of the distribution modulus, (a; whereas, for dry grinding the distribution modulus decreases slightly with increased size reduction. Wet grindability seems independent of the weight of material in the mill while dry grindability depends on weight. Experimental data and eqmtions for each process are presented for noncleavable isotropic material, for cleavable or material in which there is some form of internal stress, and for energy consumption. The equation ED/EW=BK-? (where ED and Ew are, respectively, energy for wet and dry grinding, k is the size modulus, and B and ? are constants) was found to give the ratio of energy. Dry grinding is generally considered to require more energy than wet grinding. Taggart' states that power consumption in wet grinding ranges from 60 to 90 pct of that required for dry grinding. Bond' states that dry grinding requires one third more power than wet grinding. It is interesting that Rose and Sullivan' note this same difference in power required for wet and dry grinding, but they state that on purely mechanical grounds it is difficult to see any great difference in fundamental principles between wet and dry milling, since dry milling may be regarded as wet milling with a fluid of low viscosity (i.e., air vs water). Yet it must be the viscosity of the fluid that is responsible for the differences between wet and dry grinding. Probably the most comprehensive investigation comparing wet and dry grinding has been that made by Coghill and Devaney. They compared the grind-ability of both chert and dolomite in a 19x36-in. batch ball mill under both wet and dry conditions. In these tests the weight of ore in the mill was varied from 35 to 200 lb and each grinding test was carried out for such period of time that the amount of -200-mesh material in the product was essentially the same for each test with a given material. Cog-hill and DeVaney's data for the grindability of dolomite, which are reproduced in Fig. 1, show that a comparison between wet and dry grinding is com- plex. The comminution of a material by a wet grinding method is characterized by a constant grindability over a wide range of mill loadings. Since grindability is the weight of material comminuted per unit of energy expended, it is equivalent to (w)/ (p)(t), where W is the weight of material being ground, P is the power, and t is the grinding time. In the case of wet grinding, the grindability of dolomite in the Coghill and DeVaney experiments was 0.546 * 0.009 ton per hp-hr when the load of dolomite in the mill ranged from 50 to 200 lb. On the other hand, for dry grinding no such simple relationship was found to exist between power, time, and weight of material in the mill. Dry grinding efficiency appears to depend greatly upon the amount of material within the mill. When the mill contained

Jan 1, 1961

By D. C. Shelton

Eq. 4 and 5 on page 5 of the March 1965 issue of Transactions, and Eq. 12 on page 8 of the same issue should read respectively:

Jan 1, 1965

The only technical session on Thursday was a continuation of the Symposium on Pyrometry. The members not attending the symposium -made the trip to La Salle. They were welcomed by the Mayor, to whose speech Major Arthur S. Dwight briefly responded. After luncheon the party divided and was taken by automobile to the various chosen points of interest, such as the cement works, the various zinc smelters, the coal mines, and underground cement-rock mines, where underground steam shovels were operating, etc. A special scenic trip was arranged for the ladies. The party returned by special train to Chicago, arriving at 7:30. Those desiring to visit the coal mines at Franklin and Macoupin Coun-ties left by train that evening on a special car for Benton, and were piloted through the field by Mr. Putnam, superintendent of the Old Ben Coal Corpn., Mr. Carroll, assistant general superintendent of the Chicago, Wilmington & Franklin Coal Co., and Carl Scholz, general manager of the Vallier Coal Co. Through the courtesy of the various railroads the visitors were furnished a special coach which took them to the various mines with the least delay and enabled them to see more than would be otherwise possible. Mines 8 and 9 of the Old Ben Coal Corpn., at West Frankfort, were first visited; thence the party proceeded to Orient mine of the Chicago; Wilmington & Franklin Coal Co., where luncheon was served, and then proceeded to Vallier, where both the top and the under-ground workings were inspected. The party returned to. Benton and from there proceeded in automobiles to inspect the washery plant of the United States Fuel Co. at that point; and boarded the train reaching Chicago Saturday morning. Several left Benton for their homes without returning to Chicago.

Jan 11, 1919

By Gerard Matthes

ONLY in recent years has any practical headway been made in the application of aerial photography to geological problems, and up to the present time its principal value to the geologist and mining engineer has been in "areal geology;" i. e., the study of formations as revealed at the surface. The object of this paper is to set forth briefly the principal advantages that may be derived from aerial photographic data in geological work and to serve in a measure as a guide to those who may be con-templating its use. Geophysical prospecting has for object the locating of mineral under the surface. Inasmuch as the geologist's chief complaint in life is a dual one-inability to see under the surface and inability to visualize comprehensively what he does see on top of the ground-geophysical prospecting coupled with aerial photography should do much to lessen his burdens. To what extent aerial photography can be made a useful adjunct to geophysical prospecting remains to be proved, but the inference drawn here is that it holds many alluring possibilities. To avoid disappointments, it should be stated at the outset that to the :geologist or mining engineer an aerial photograph or mosaic is only a tool, the effecient use of which he must set out to master. Upon his skill and ingenuity in applying it to particular problems will depend largely the value of the results that he will obtain. Under the most favorable conditions an aerial mosaic or even a few loose photographs, because of the exceptionally comprehensive oversight which they afford of field data, such as dips and strike observations at isolated outcrops, local indications of faulting, mineral and fossil finds, greatly facilitate study and correlation, save much tedious foot work, and may repay their cost many times. Even so, types of country may be found where the most, skillful use of the best of pictures will produce results so meager as to render the cost of aerial photography entirely unwarranted.

Jan 1, 1928

By E. I. Alessandrini, A. U. Seybolt

Selective oxidation (preferential oxidation of a single alloy component) can be an important oxidation mode in the case of alloys such as Si-Fe, Al-Fe, and so on, when heat-treating in atmospheres of low oxygen potential. It was the purpose of this work to identify the oxide films formed on iron containing 31/4 pct Si at low oxygen potentials, and to make some observations on the H2O/H2 ratios required to form them. EXPERIMENTAL METHOD HYDROGEN of varying water content offers a very convenient method of establishing a wide range of oxidizing power, or oxygen potential. The method chosen to provide a desired H2O/H2 ratio was to nearly saturate hydrogen at room temperature by bubbling the gas through water at room temperature, and then to wring out the excess water by passing the wet gas through a long glass spiral immersed in a refrigerated liquid at a desired dew point. After oxidation, the samples were examined by electron diffraction in order to identify any film which might be present. PREPARATION AND TREATMENT OF SAMPLES The samples were prepared from a Si-Fe alloy made in the laboratory using electrolytic iron and 96 pct ferro-silico11, vacuum-induction melted, hot-rolled then cold-rolled to 0.015-in. thickness. After the cold-rolling the samples were annealed at 1000°C. Transmission Laue X-ray patterns showed very little evidence of preferred orientation, but it is likely that a weak texture could have escaped notice. The chemical analysis was as follows: 3.34 Si, 0.01 Al, 0.001 0, 0.006 C, 0.002 N. Before each sample was oxidized, it was polished through 4/0 emery paper, cleaned and dried using benzene, carbon tetrachloride, acetone, absolute alcohol, and ether. The method of oxidizing the samples was as follows: Samples about 1/2 by 1 by 0.015 in. were held in the upper, cool part of a vertical vacuum furnace until the desired temperature was established in the furnace, Fig. 1. The sample was then lowered into the uniform hot zone where it was heated quickly to the desired temperature, and subjected to hydrogen of controlled moisture content (dew point) as mentioned previously. Temperatures from 400" to 1000°C were used, and dew points from room temperature to —78°C were explored. In general, the runs lasted about 6 hr, but a few runs were of different duration, as will be indicated later. After the predetermined period of time had elapsed, the specimens were withdrawn by the hollow stainless steel-quartz holder to a cool zone at the top of the furnace. EXPERIMENTAL RESULTS The results are summarized in Table I. Most of the electron-diffraction patterns were obtained by glancing-angle reflection, but in a few cases it was possible to strip the oxide films from the surface by a total immersion in a 3 pct bromine-methanol solution. Some of the thinner stripped films yielded quite satisfactory transmission patterns. However, both diffraction techniques yielded the same results; the patterns while most frequently showing a (low) tridymite, also showed a (low) cristobalite in some cases. A few patterns showed evidence of both phases. Because of the known nonequilibrium behavior of silica, it is not surprising that both forms of silica could coexist on one sample. It was noted that at fairly low dew points such as —40°C, there was a tendency for diffuse diffraction rings, suggesting a very small crystallite size. Also some subsequent experiments to be reported later show that thin silica films can be amorphous.

Jan 1, 1959

By Castle. J. E., J. L. Gillson

In this chapter there is a wide difference in the meaning of some words used for rock and mineral names as defined by scientists and tabulated in the "Glossary of Geology and Related Sciences," published in 1957 by the American Geological Institute, and the meaning as used by the commercial producers of feldspar. It is necessary, therefore, to define these terms giving both the "scientific" and the "practical" definitions. There is no ambiguity about the name feldspar itself. This is a general term for a group of related minerals which are silicates of aluminum, combined with sodium, potassium and calcium. The feldspars are the most abundant minerals of the igneous rocks, and since these rocks make up the largest proportion of the earth's crust by volume (although not of the actual surface area of the continental land masses) the feldspars are the most abundant minerals of all. The igneous rocks themselves are classified on the basis of the variety of the feldspar present. The potash feldspars, which are called orthoclase and microcline, are the dominant minerals of the granites and syenites, and of their fine grained equivalents, the rhyolites and trachites. The plagioclase feldspars form an isomorphic series, of which the sodic endmember is called albite, and the lime endmember, anorthite. The members near the sodic end, albite, oligoclase and andesine are the characteristic feldspars of the granodiorites and diorites, and of their fine grained equivalents, the dacites, and andesites. These feldspars-orthoclase, microcline, albite oligoclase and andesine-are the feldspars used commercially. Labradorite, a member of the plagioclase group, in which the lime exceeds the soda in the chemical composition, is the common feldspar of the gabbros, anorthosites, norites, basalts and other so-called "basic" rocks. Although some attempts have been made to use labradorite as a commercial feldspar, none is now used except as an ornamental stone. The mineral anorthite is very rare. Since the feldspar in most of the common igneous rocks is relatively fine grained, and mixed intimately with other minerals, from which it can be separated cleanly only with difficulty, most masses of the common rocks cannot be used as sources for the feldspar contained in them. It is only the rarer types of rocks in which the feldspar is very coarse-grained, or ones containing very little of the iron-bearing minerals which are objectionable impurities in commercial feldspar, that are quarried to yield an acceptable product. The most widespread type of such a rock from which commercial feldspar can be produced, is that called "pegmatite." These rocks are dikes and bodies of irregular shape and of a size measured in feet or yards in most cases, rather than in miles. They are more commonly associated with granites, rather than with other coarse-grained igneous rocks. Many of them contain crystals of various minerals, including the feldspars, that are much larger than are found in the host rock, the granite, and in many there are large masses of feldspar that are sufficiently free from iron-bearing silicates so that products acceptable to the ceramic industry can be produced, in some cases simply by hand sorting. Most of this feldspar in pegmatites, however, does contain intergrown quartz up to 25 pct. Pegmatites are so important, not only as commercial sources of feldspar, but also of other minerals which are produced com-

Jan 1, 1960

By Michael Epstein, Daniel E. Rosner

Fume nucleation sufficiently close to vaporizing suvfaces can augment net vaporization rates into cooler environments. Environmental conditions favoring large vaporization rate enhancements are briefly discussed and a previous theoretical treatment of this nucleation phenomenon is generalized to account for the self-regulating effect of condensalion-heat release within the boundary layer. Despite kinetic limitations on homogeneous nucleation, and latent heat release, non-diffusive condensate removal processes appear to make possible large enhancements in steady-state vaporization rates, provided surface temperatures are well below the boiling point. When condensed phases vaporize (or dissolve) into cooler media, the diffusion-limited mass loss rate can be strongly influenced by the process of nucleation/con-densation (or precipitation) within the thermal boundary layer. This condensation process (which typically leads to mists or fumes in the case of evaporation into cooler gases) has the effect of steepening the vapor pressure profiles near the evaporating surface, since the condensation zone acts as a vapor sink. However. the resulting enhancement in the diffusion-limited evaporation rate can be estimated (as first done by Turkdogan1 for the case of molten iron/nickel alloys evaporating into helium) only if one has independent knowledge of the critical supersaturation, sCrit(T), required to homogeneously nucleate the vapor.* In a recent reformulation and generalization of the theoretical model of Ref. 1 it has been shown that, when log sCrit is approximately linear in reciprocal temperature, rather simple expressions can be derived4 for the ratio of the actual rate of vaporization j" to either the minimum (no condensation) rate j"min, or the maximum (equilibrium condensation) rate j"max In the present communication we wish to briefly report on further developments and implications of the formulation of Ref. 4, with emphasis on i) environmental conditions favoring large enhancements in vaporization rate, and ii) the self-regulatory influence of condensation heat release (neglected in Refs. 1 to 4) on predicted vaporization rates. Additionally, we take this opportunity to correct several misprints appear- ing in Ref. 4, and comment on Elenbaas's recent criticism5 of Ref. 1. MAXIMUM POSSIBLE VAPORIZATION RATE IN PRESENCE OF CONDENSATION A nonequilibrium theory is of interest because of the very large difference between the minimum (no condensation) and maximum (equilibrium condensation) vaporization rate. The magnitude of this maximum possible enhancement can be shown quite clearly by combining a result of Refs. 3 and 4 with the fact that for most liquids there is a simple relationship between the molar heat of evaporation, A, and its boiling point, i.e., A/(RTBp) = C, where the constant C, often called the Trouton ratio, takes on values not very different from 11.* More generally, for any substance (including The Trouton ratio (which for water is 13, for methane, 10, and so forth) will be recognized as the ratio of the molar entropy change upon vaporization (at TBP or Ttransf) to the unlversal gas constant R. Its near constancy reflects the fact that the change in atomic order upon vaporization depends only weakly on the kinds of molecules involved. those that sublime under ordinary conditions) we can define a characteristic transformation temperature. Ttransf, by a relation of the form Ttransf =A/(CR), and then examine the maximum possible evaporation rates as a function of how far removed from Ttransf are the surface temperature, Tw, and ambient temperature, T. Subject to the assumptions: 1) equilibrium vapor pressure, pv,eq, everywhere small compared to prevailing total pressure, p, and 2) negligible effect of condensation heat on temperature profile, the maximum enhancement ratio was found (Eq. [17], Ref. 4) to be: where, for most vapors, Nu/NuD (the ratio of heat transfer coefficient to mass transfer coefficient for the same configuration) is a number near unity.* Ex- *An alternative derivation of the Nu = NuD special case of this equation. revealing its validity for arbitrary velocity/temperature profiles in a laminar boundary layer, is given in Ref. 3. amining this result for a "Trouton substance", one obtains the results shown in Fig. 1, constructed for C = 11. Since we are concerned with vaporization enhancements (j'max/J"min > 1) at surface temperatures below Ttransf, this area of interest is shown unshaded. One notes that at a fixed ambient temperature (hence, T/TtranSf) there is a unique surface temperature, 2T , at which j"max/j"min attains its peak value; moreover, the peak enhancement ratio, see dashed locus. Fig. 1, is: (NuA/NuD)(C/4)(Ttransf/T,). Hence, if Nu = NuD, when the ambient temperature is less than 1/4 of TtranSf the peak enhancement exceeds the Trouton

Jan 1, 1969

By Charles L. Mantell, James E. Hoffmann

Mechanical inclusion, electrophoretic deposition, and adsorption were studied as mechanisms for code-position of aluminas present in copper-plating electrolytes as an insoluble disperse phase. Mechanical inclusion was not a significant factor. That codeposi-tzon of aluminas by an electrophoretic mechanism was unlikely was substantiated by measurements of the potential of the aluminas. The alumina content of the deposits was studied as a function of the pH of the bath. These tests in conjunction with sedimentation studies demonstrated the absence of an isoelectric point for the alutninas over the pH range examined. Thiourea in the electrolyte (a substance known to be adsorbed on a copper cathode during electrodeposition) affected the amount of alumina in the electrodeposit. However, no adsorption of thiourea on aluminas in aqueous dispersions was detected. If it were possible to produce a dispersion-hardened alloy of copper and alumina by electrodeposition, an alloy possessing both strength and high conductivity at elevated temperatures might be anticipated. Investigation of the mechanism of codeposition of aluminas with copper was undertaken with the hope that knowledge of the mechanism would aid in the development of such an alloy. The word "codeposit" here does not necessarily imply an electrolytic phenomenon but rather that the materials codepositing, the various aluminas, are transported to and embedded in the electrodeposited copper by some means. Mechanical inclusion in electrodeposition implies a mechanism of codeposition which is wholly mechanical in nature; the only forces acting on a particle are gravity and contact forces. Such a particle is presumed to be electrically inert and incapable of any electrical interaction with electrodes in an electrolytic plating bath. Processes for matrices containing a codeposited phase by electrodeposition from a bath containing a disperse insoluble phase frequently state that code-position is caused by mechanical inclusion.10,2,12 If settling, i.e., gravity, be the controlling mechanism for codeposition of aluminas, then assumptions may be made that 1) the content of alumina in the electrodeposit should be enhanced by increasing the particle size, 2) the geometry of the system, that is, the disposition of the cathode surfaces relative to the di- rection of the falling particles, should affect the alumina content of the electrodeposit, 3) in geometrically identical systems the chemical composition of the electrolyte employed should exercise no effect on the alumina content of the deposit, that is, the alumina content should be the same in all cathode deposits irrespective of bath composition. A bent cathode19 evaluates the clarity of filter effluent in electroplating baths by comparing the roughness of the deposit on the vertical surface with that on the horizontal surface. Two difficulties are inherent in this technique: 1) the current density on the horizontal portion of the cathode would be substantially greater than that on the vertical surface; 2) should the deposit obtained be rough, projections on the vertical face could act as horizontal planes and vitiate the relationship between the vertical and horizontal surfaces. Bath composition should have no substantial effect on the alumina content of the deposit. Two different electrolytic baths were employed. They possessed variant specific conductances and substantially different pH ranges. The experimental tanks were rectangular Pyrex battery jars 6 in. wide by 3 1/4 in. long by 9 3/4 in. deep. The cathodes were stainless steel 316 sheet of 0.030 in. thickness, cut to 7.5 by 1.75 in. and bent at right angles to form an L-shaped cathode whose horizontal surfaces measured 1.75 by 3.0 in. All edges and vertical surfaces were masked with Scotch Elec-troplaters Tape No. 470. The anodes were electrolytic cathode copper 9 in. high by 2.25 in. wide by 0.5 in. thick. To eliminate inordinately high current densities on the projecting edge of the cathode, the anode was masked 1 in. above and below the projected line of intersection of the cathode with the anode. The exposed area of the anode was equal to that of the cathode, providing both with equal average current densities. The agitator in the cell was of Pyrex glass and positioned so its center line was equidistant from cathode and anode, and a plane passed horizontally through the center of the blade would be located equidistant from the bottom of the cathode and the bottom of the deposition tank. The assembled apparatus is depicted in Fig. 1. Hatched areas on anode and cathode represent the area of the electrodes wrapped with electroplaters tape. MATERIALS The chemicals were copper sulfate—CuSO4 • 5H2O— technical powder (Fisher Scientific Co.). Spectro-graphic analysis showed substantial freedom from antimony, arsenic, and iron. Traces of nickel were present.

Jan 1, 1967

By I. R. Kramer

Studies of the effect of size of the specimen on the change of slopes of Stages I and 11 by surface removal showed that the change of Stage I was independent of size with respect to the polishing rate; however, the change in the slope of Stage 11 with polishing rate increased directly in proportion to the surface area. The removal of the surface during the test affected the plastic deformation characteristics of gold, aluminum, and zinc single crystals and polycrystalline aluminum. The apparent activation energy of aluminum was found to be decreased markedly by removing the surface during the deformation process. In previous papers1-3 it was shown that the surface played an important role in the plastic deformation of metals. By removing the surface layers of a crystal of aluminum by electrolytic polishing during tensile deformation, it was found that the slopes of Stages I, II, and III were decreased and the extents of Stages I and II were increased when the rate of metal removal was increased. By removing a sufficient amount of the surface layer after a specimen had been deformed into the Stage I region, upon reloading, the flow stress was the same as the original critical resolved shear stress and the extent of Stage I was the same as if the specimen had not been deformed previously. The slope of Stage I was decreased 50 pct and that of Stage 11 decreased 25 pct when the rate of metal removal was 50 X 10"5 ipm. These data show that in Stage I the work hardening is controlled almost entirely by the surface conditions, while in Stages 11 and III both surface conditions and internal obstacles to dislocation motion are important. It appears that during the egress of dislocations from the crystal, a fraction of them becomes stuck or trapped in the surface regions and a layer of a high dislocation concentration is formed. This layer would not only impede the motion of dislocations, but would provide a barrier against which dislocations may pile up. In this case, there will be a stress, opposite to that of the applied stress, imposed on the dislocation source and dislocations moving in the region beyond this layer. It has been found convenient to refer to this layer as a "debris" layer. The "debris" layer may be similar to the dislocation tangle observed by thin-film electron microscope techniques.4 Reported in this paper are the results of studies on the effects of removing the surface during plastic deformation on aluminum crystals of various sizes. The effects of the surface on the yield point behavior of gold and high-purity aluminum crystals as well as the creep behavior were also determined. The effects of surface removal on polycrystalline aluminum (1100-0 and 7075-T6) are also reported. EXPERIMENTAL PROCEDURE For those portions of the investigation involving creep and tensile specimens, single crystals, having a 3-in. gage length and a nominal 1/8-in. sq cross section, were prepared by a modified Bridgman technique using a multiple-cavity graphite mold. The single crystals were prepared from materials which had initial purities of 99.997, 99.999, 99.999, and 99.999 pct for Al, Cu, Zn, and Au, respectively. The aluminum specimens for the size effect studies were prepared through the use of a three-tier mold in which crystals having a cross section of 1/8, 1/4, and 1/2 in. were grown from a common seed. The mold design was arranged so that one 1/2-in. crystal, two 1/4-in. crystals, and four 1/8-in, crystals of the same orientation could be cast. With this technique, it was possible to obtain only one set of crystals with the same orientation. Because of this limitation, it was not possible to determine both the changes of extent and slope of the various stages since a large number of crystals of the same orientation would have been required. Instead, only the change of slope as a function of the rate of metal removal was studied by abruptly altering the current density of the electrolytic polishing bath at various strains within the regions of Stages I and 11. The experimental techniques used for the tensile studies were essentially the same as those used previously.1,3 The specimens were deformed in a 200-lb Instron tensile machine, usually at a rate of 10-5 sec-5. A methyl alcohol-nitric acid solution was used as the polishing bath for aluminum. The temperature was maintained constant within ±0.l°C by means of a water bath. The tensile machine was

Jan 1, 1963

By R. W. Bartlett

The rates of oxidation of tungsten have been determined at temperatures between 1320" and 3170°C and oxygen pressures to 1 amn using a surface -recession measurement technique. Above approximately 2000°C and 10-6 atm the rate is independent of temperature and can be calculated from gas collision theory assuming a constant reaction probability, e, of 0.06. Oxygen molecules react at surface sites where oxygen atoms have previously chemisorbed. This provides a direct pressure dependence at low pressures but at high pressures tungsten oxide molecule s form an adjacent gas boundary layer which lowers the PO2 at the tungsten surface. A correction for this effect using free-convection theory fits the rate data over the entire oxygen-pressure range from 10-8 to 1 atrn as well as data using O2-A mixtures. Below 10-6 atrn and above 2000°C, e decreases with increasing temperature because of desorption of oxygen atoms. Below 2000°C the rate decreases with decreasing temperature at all oxygen pressures following an apparent activation energy of 42 kcal per mole and depending on (Po2)n with n varying between 0.55 and 0.80. MOST of the previous tungsten oxidation studies have employed gravimetric methods and have been limited to temperatures below 1000°C where the weight loss associated with evaporation of tungsten oxides is negligible compared with the weight gain from oxidation.' At higher temperatures, oxygen-consumption rates have been determined from pressure measurements, usually at constant flow rates, by Langmuir,2 Eisinger,3 Becker, Becker, and Brandes,4 and Anderson.5 The sensitivity of this method decreases with increasing pressure and, with the exception of Langmuir's work, these investigations were confined to pressures below 10-6 atm. Above approximately 1300°C, depending on the oxygen pressure, the rate of oxide evaporation is greater than the oxide-formation rate and the recession of the tungsten surface can be measured optically without interference from an oxide layer. This was first done by Perkins and crooks6 who heated tungsten rods in air pressures from 1 to 40 torr at temperatures between 1300" and 3000°C. The present investigation of the oxidation kinetics of tungsten at high temperatures emphasizes oxygen pressures from 10-6 to 1 atm. This is the range of interest for earth atmosphere re-entry applications of tungsten for which little data were previously available. APPARATUS The apparatus is a modification of the type used by Perkins and crooks.' Ground tungsten seal rods, 6 in. long by 0.125 in. diam, were mounted vertically between two water-cooled electrodes, one fixed and the other having free vertical travel. The movable counter-weighted electrode is prevented from undergoing horizontal displacement by three sets of runners mounted at 120-deg intervals. Electrical contact is made by means of a water-cooled mercury pool. A 24-in. vacuum bell jar having a volume of approximately 267 liters was used as the reaction chamber with the sample holder mounted in the middle of the chamber. Power was supplied from an 800-amp dc variable power supply. Temperature readings were made by means of a Latronics automatic two-color recording pyrometer. With this instrument, corrections for emissivity are not necessary provided the spectral emissivi-ties at two closely spaced wavelengths are equal. Supporting measurements were made with a micro-optical pyrometer corrected for emissivity of bare tungsten and window absorptivity. The micro-optical pyrometer was calibrated against a National Bureau of Standards calibrated tungsten lamp and both pyrometers were periodically checked against the melting points of tungsten and molybdenum using the oxidation apparatus. Above 10-6 atm, pressures were measured with an Alphatron gage calibrated against a McCleod gage. At 10-6 atm, a hot-filament ionization gage was employed. A magnified image of the self-illuminated tungsten rod was formed using a 360-mm objective lens mounted outside the bell jar. When the experiment exceeded 1 hr, the image was focused on a ground-glass plate about 10 ft from the tungsten rod at about X8 and the recession of the thickness of this image was monitored with a Gaertner cathe-tometer. When faster rates were encountered, a 35-mm time-lapse cinecamera with a telephoto lens and bellows extension was substituted for the ground-glass plate and cathetometer. Diameter recession rates were determined from the photograph image projected on the screen of an analytical film reader. EXPERIMENTAL PROCEDURE After installing the rod in the apparatus and cleaning it with acetone, the system was evacuated to 5 1 x 10-5 torr. Before oxygen was introduced,

Jan 1, 1964

By Marius S. Darrow, William B. White, Rustum Roy

The PbS-PbTe systen has been studied by quench-ing and D.T.A. techniques f?om 400' to 1150°C. Runs were made in evacuated silica tubes so that all equilibria are at the vapor pressure of the system. Lattice parameters of the quenched salnples , measured by X-ray diffraction, show a complete crystalline-solution series existing over a narrow temperature range between approximately 805" and 871°C. An exsolution dome extends from a maximum of about 805"C (approximately 30 mole pct PbTe) to 1 and 96.5 pet PbTe at 400°C. A narrow melting region, deternined by D.T.A., extends form 918c (mp PbTe), The shapes of the liquides and solidus curves imply the existence of a minimum at 871°C at approximately 65 pct PbTe. THe exact composition of the minimum could not be established due to the very narrow two-phase region. At compositions containing less than 50 pet PbTe, liquidus temperatures begin to increase, while the solidus remains almost flat to about 15 mole pet PbTe before beginning to vise toward the mp of PbS (1075 C). LEAD sulfide and lead telluride are isostructural (NaC1 type) semiconductors whose electrical and optical properties have been extensively studied and used in recent years. If appreciable crystalline solution exists between these compounds, the variation of physical properties with composition could be of interest. The purpose of this investigation was to determine the extent, if any. of crystalline solution, and to obtain the phase diagram for the system. To the knowledge of the authors, only three studies of the system PbS-PbTe have been reported, and, in chronological order, each investigation found an increasing amount of crystalline solution. In 1956, Yamamoto reported finding no evidence of crystalline solution between the compounds. Sindeyeva and Godov-ikov,' in 1959, found very limited crystalline solution. but only under conditions of excess tellurium concentration. Finally Melevski s3 investigation in 1963 indicated that one solid phase exists in the region from PbS to 7 pct PbTe and from 82 pct PbTe to PbTe at 886'C, with an eutectic at 55 pct PbTe at that temperature. Detailed data on the solvus boundary were not given. EXPERIMENTAL EQUIPMENT AND MATERIALS Commercially produced PbTe and PbS powders were used as starting materials. Batches of specific mole percent composition were accurately weighed and mixed in a plastic bottle, in a shaker mill. An analy- sis of impurity content is given in Table I for pure PbS and PbTe and for two randomly selected batches after the powders were mixed. Individual samples, ranging in weight from 0.2 to 0.5 g, were sealed in evacuated silica tubes which had been thoroughly washed and rinsed with acetone and distilled water. Thus all data taken were at the pressure of the system. Subsolidus relations were studied down to 400°C by heating the samples in a vertical tube furnace for 24 hr. The sealed tubes were quenched in water with quench time from the hot zone not exceeding 1 sec. Temperatures were measured by a chromel-alumel thermocouple and controlled to 53°C for most runs. The number and composition of phases present were determined from powder X-ray diffraction patterns taken at room temperature on a Norelco diffractome-ter, using silicon as an external standard. Above 850°C quenching techniques were, in general, found to be unsatisfactory, and differential thermal analysis (D.T.A.) was used to determine melting relations. The evacuated tubes were recessed about 1 cm at one end to accommodate the differential thermocouple. Al203 was used as the reference material in a similar tube containing the other side of the differential couple. For temperature measurements, a separate thermocouple was placed in the recess of the tube containing the sample to be measured, thus providing an opportunity to obtain thermal, as well as differential, analysis. All thermocouples for these measurements were Pt-Pt 10 pct Rh. Temperature and differential curves were recorded separately on synchronized strip-chart recorders. Thermocouples and recording equipment were calibrated using NaCl and gold standards, using the melting points 801" and 1063 C, respectively, which span most of the temperature range of interest. Heating and cooling rates generally were from 4 to 7°C per min. It was found, in fact. that rates ranging from 1.5 to 25°C per min did not significantly change the data obtained.

Jan 1, 1967

By L. F. Mondolfo, W. T. Collins

A study of the relationship between undercooling for nucleation and structure in Sn-Cu alloys with 0.1 to 5 pct Cu has shown that in hypereutectic allojls the halo of tin that surrounds the primary crystals of Cu3Sn5 is larger, the larger the undercooling for nucleation o,f the tin. This increase of halo size results in a decrease of coupled eutectic, and, in alloys far from the eulectic composition, may produce its complete disappeavance, with the formation of a divorced eutectic structure. This was confirnred by the excrrnination of other alloys in which divorced eutectic slructuves are formed, and leads to the conclusion that they ave only an extrenle case of halo forrtzalion , which results when the two phases freeze one at a time and solidification of the first is completed Defove the second starts. It was also found that under proper conditions of nucleation all types of eutectic structures can be formed in the sartte system , and therefore divorced eutectics, like normal and anomalous, are not characteristic of the syslett~, but are mainly controlled by nucleatiorz. Dizlovced eutectics are formed when the phase that tutcleates the eulectic vequires a large undevcooling for ils nucleation and when the cotnpositiorz of the alloy is far from the eutectic., on the side of the primary phase that does not nucleate the other phase. It is recommended that the tevm "divorced" be used in preference to degenerate because it is more desct-iptice of the morphology and mode of forinalion of the structures. ThE variety of structures found in eutectic alloys has been extensively investigated and classified. The most accepted classification is the one by ~cheil,' in which three different types of eutectic were distinguished: 1) normal, 2) anomalous, 3) degenerate (divorced). ATornlal eutectics are typified by the simultaneous growth of the two phases ("coupling") by which the two phases appear as interpenetrating crystals. The presence of a crystallization front, in which the two phases grow side by side, creates the eutectic grains, with the boundaries where the fronts meet. The presence of eutectic grains is the .distinguishing feature of a normal eutectic, according to Scheil. Straumanis and Brakss2 examined the Cd-Zn system and showed that there was a crystallographic relationship between the phases. Later, others4 also investigated additional systems and found definite crystallographic relationships in the coupled eutectics. The anornalous eutectic shows much less coupling than the normal; the two phases are intimately mixed but 'grow without a uniform crystallization front—a consistent crystallographic relationship— and the eutectic grain is conspicuously absent. As in the normal eutectics faster rates of growth result in a finer structure, but there is not the typical uniform spacing of normal eutectics. The degenerate eutectic shows no coupling; in fact the two phases attempt to minimize their area of contact and to form separate crystals. It has been suggested5" that slow cooling may favor this type of structure. Scheil believes that normal eutectics are formed when the two solid phases are present in more or less equal proportions, whereas both anomalous and degenerate eutectics form when one of the phases is present only in small amounts. spengler7 extended much farther this qualitative relationship between the eutectic type and the ratio of the two phases, and added a relationship to the melting point of the constituents. On this basis he proposed two equations for determining into which of Scheil's classifications an alloy belongs. The first equation is: where TI is the melting temperature of the lower-melting component, Tp of the higher-melting component, and Te the eutectic temperature. The second equations is: where is the volume percent of the lower-melting phase and $2 of the higher-melting phase at the eutectic composition. If 0 and/or 4 are in the range 0.1 to 1, a normal eutectic is formed; if in the range 0.01 to 0.1, anomalous; if less than 0.01, degenerate. Although the examples given by Spengler show a good agreement with the formulas, chadwick found that the Zn-Sn eutectic is normal to all growth rates, even though the volume ratio is 12/1, and Davies9 reports that the A1-AlgCo2 eutectic is normal, with a volume ratio of more than 30/1. Many more discrepancies of this type can also be found. Neither Scheil nor most of the other investigators have considered nucleation as a factor in the formation of divorced eutectics. Daviesg states that divorced eutectics form when neither phase acts as

Jan 1, 1965

By David A. Thomas, David N. French

The lrnrdness of WC crystals has been measured with the Knoop indenter at loads of 100 and 500 g on the (0001) and (1070) planes. The hardness as tneasitred on the basal plane is 2400 kg per sq mm and shows little anisotropy. The hardness on the prism plane, however, shows a marked orientation dependence, with a low value of 1000 kg -per sq mm when the long axis of the Knoop indenter is oriented parallel to the c axis and a high value of 2400 kg per sq mm when the indenter is perpendicular to the c axis. Slip lines (Ire observed surrounding the microhardness indentations and they show slip on (1010) planes, probably in [0001] and (1120) directions. This slip behavior can be explained by the crystal structure of TVC, which is simple hexagonal with a c/a ralio of 0.976. The hardness anisotropy call be explained by [0001]{1010} and (1130) {10l0] slii) and the resolved shear-stress analysis of Daniels and Dunn. HARDNESS anisotropy is a well-known phenomenon and has been reported for many metals, with both cubic and hexagonal structure.1-6 For hexagonal tungsten carbide, WC, a wide range of hardness values is reported in the literature. For example, Schwarzkopf and Kieffer7 give a hardness of 2400 kg per sq mm and report a value of 2500 kg per sq mm by Hinnüber. Foster and coworkerss give the average Knoop microhardness as 1307 kg per sq mm with a maximum value of 2105 kg per sq mm. Although these values and the structure of WC suggest the likelihood of hardness anisotropy, no such measurements have been made. We first detected a large apparent hardness anisotropy in WC crystals about 75 p large, in over-sintered cemented tungsten carbide. Prominent slip lines also occurred around many indentations. This report presents further observations and interpretations of hardness anisotropy and slip in WC crystals obtained from Kennametal, Inc. Both Knoop and diamond pyramid indenters were used on a Wilson microhardness tester with loads of 100 and 500 g. EXPERIMENTAL RESULTS The carbide crystals tended to be triangular plates parallel to the (0001) basal plane of the hexagonal structure. The side faces were parallel to the ( 1010) prism planes. Specimens were mounted approximately parallel to these two types of faces and metallographically polished. Laue back-reflection X-ray patterns were used to orient the specimens, which werethen ground to within ±1 deg of the (0001) and (1010) planes. The Knoop hardness values measured on the basal plane are plotted in Fig. 1. There is only a small anisotropy, with a hardness range of 2240 to 2510 kg per sq mm. The additional points at angles from 52.5 to 67.5 deg confirm the sharp minimum hardness at 60-deg intervals, consistent with the sixfold hexagonal symmetry. The average hardness of all values obtained on the basal plane is 2400 kg per sq mm. While the basal plane shows only slight anisotropy, the (1010) plane exhibits marked hardness anisotropy, from 1000 to 2400 kg per sq mm. Fig. 2 shows the hardness as a function of the angle between the long axis of the indenter and the hexagonal c axis, the [0001] direction. The minimum and maximum occur when the indenter is oriented parallel and perpendicular to the [0001] direction, respectively. The anisotropy of the prism plane is contrary to that reported for hexagonal zinc and hard- However, the basal-plane anisotropy is similar to these two metals.1'2 To check the accuracy and reproducibility of the measurements, a series of fifteen impressions was made at 100-g load in the same orientation in the same area of the specimen surface. The average for all was 2040 kg per sq mm, with a range of 1950 to 2130 kg per sq mm, giving an accuracy of about ± 5 pct. Thus the slight anisotropy on the basal plane is almost within experimental error. Fig. 3 shows slip lines around the Knoop indentations on the basal plane. The slip traces are in directions of the type (1130). The presence of slip steps on the basal plane indicates that the slip direction lies out of the (0001) plane. Because WC has a c/a ratio of 0.976,' the shortest slip vector is [0001], which suggests slip systems of the type [0001] (1010). Fig. 4 shows slip lines around the Knoop intentations on the (1010) plane. These slip lines are inconsistent with [0001] slip but can be

Jan 1, 1965

By D. L. Kelley

High-viscosity water injection has been proposed for use in reservoirs containing high-viscosity crude oils. Previous publications have largely ignored the possible effects of the connate water on the proposed process. This paper describes experimental work which indicates that the connate water will be forced ahead of the injected water to form a bank of low-viscosity water. This decreases the oil recovery which would be expected if such a bank were not formed. These effects are shown for a range of fluid mobilities and connate-water saturations for a five-spot injection system. In general, oil recoveries using viscous water are significantly greater than for untreated water even though they are less than would be expected if no connate water bank were formed. INTRODUCTION The effect of mobility ratio on the oil recovery of wa-terfloods has been known for many years. Muskat first pointed out that the fluid mobilities (k/µ) in the oil and water regions would affect the performance of the water-flood, and he estimated the general effect of these variables.' Since this early work, studies of the effect of mobility ratio on secondary recovery have been reported where mathematical,' potentiometric3 and scaled flow models' were used. These studies have shown that a reduction in the mobility ratio between the oil and the displacing fluid would cause additional oil recovery when water-flooding reservoirs containing viscous crude oils. Studies reported by Pye- nd Sandiford 8 have indicated that chemicals to increase injection water viscosity are now available and can be used to reduce the over-all mobility ratio of a waterflood. Where mobility ratios are controlled by the injection of viscous fluids, the connate water of the reservoir can play an important part in the displacement of the reservoir oil. The purpose of this study was to determine the effect of the connate-water saturation in waterfloods where viscous waters are used for injection. DISPLACEMENT OF THE CONNATE WATER Russell, Morgan and Muskat7 were the first to recognize the mobility of connate waters in waterflooding. They conducted waterfloods on oil-saturated cores containing 20 and 35 per cent irreducible water saturations, and found that from 80 to 90 per cent of the "irreducible" water was produced after only one pore volume of water was injected. However, their experiments were conducted at rates of flow significantly higher than those ordinarily occurring in waterfloods. Also, the cores were only from 4.0 to 8.5 cm long. Brown 4 studied a 100-cm linear sand pack which had been prepared to contain connate water and oil. He used 140- and 1.8-cp oils with injection water of essentially the same viscosity as the connate water. He found that all of the connate water was displaced by the injection water in both cases. However, the injection volumes required for complete displacement of the connate water were considerably higher in the case of the more viscous oil. To verify the results of the foregoing experiment, a 10-ft-long linear model was constructed by packing 250-300 mesh sand in a 1/2-in. diameter nylon tube. The model was evacuated, saturated with a brine of 1-cp viscosity, and flooded with a 41-cp mineral oil to the irreducible water saturation of 10.9 per cent. The model was then waterflooded by the injection of a water solution which had an apparent viscosity of 42.6 cp. The solution consisted of 0.5 per cent methylcellulose in distilled water. The viscosities of the oil and connate water were measured with an Ostwald viscosimeter. The viscosity of the polymer solution was calculated by Darcy's law using pressures measured during actual flow conditions. The ratio of the mobility in the oil region to the mobility in the inject ion-water region was approximately 0.32. The mobility ratio of the oil region to the connate-water bank was approximately 14. The mobility ratio between the connate-water bank and the injection water region was 0.024. Approximately 84.5 per cent of the recoverable oil was produced before water breakthrough. Immediately following breakthrough, oil and connate water were produced at an increasing water-oil ratio until the viscous injection water broke through. At viscous-water breakthrough, 96 per cent of the original connate water had been produced. After breakthrough of the viscous water, there was no additional production of connate water or oil. The near-

Jan 1, 1967

By B. C. Riblet

THE president of a large smelting and refining company once telegraphed me that they had acquired extensive mining interests in northern Peru and requested me to inspect their properties as it would be necessary to install a system of aerial tramways between their mines and treatment plant. The concentrator was about 50 mi. from the coast, reached by a fairly good motor road. From the concentrator an aerial tramway would he required, about 6 mi. long, to their first property, a gold and silver mine. The lower terminal of the tramway was at an elevation of 4000 ft. and the mine at an elevation of 9500 ft.

Jan 1, 1937