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By E. M. Wise, T. E. Kihlgren, N. B. Pilling

Systematic data are presented on the relation of composition of nickel silvers to color, tarnish resistance, hardness and liquidus temperatures, for alloys containing up to 30 per cent nickel and 50 per cent zinc, further modified at certain nickel and zinc contents by 0 to 8 per cent of tin and 0 to 10 per cent of lead. Attention has also been directed, using a still more limited range of alloys, toward a study of tensile properties and such casting properties and characteristics as fluidity, shrinkage and pressure tightness. In some of these, rather broad ranges of composition are dealt with; in others, it has been expedient to focus attention on the Federal Specification 20 per cent nickel silver casting alloy, WWP-541. Introduction The demand for an alloy essentially white in color, which could be cast and worked by simple methods and which possessed strength and corrosion-resisting properties superior to the usual brasses, led to the discovery and early use of cupronickel alloys containing zinc, and sometimes tin, lead and other metals. This type of alloy was generally known as 'German silver until the war, when the term "Nickel Silver" became generally accepted. The element principally responsible for the white color of the alloy and for its strength and corrosion resistance is nickel, which may constitute from 8 to 30 per cent of the alloy. The functions of the other elements will be described later. The manifold uses of nickel silver are largely due to its color, corrosion resistance, ease of casting, fabrication, joining and rather moderate cost. In many instances the alloy is finished by grinding and polishing while in others it may be finished with a silver or nickel plus chromium plate. Electroplates applied to a nickel silver base behave more satisfactorily than those applied to the softer and less resistant alloys, and owing to the white color of the underlying metal the article will retain a pleasing appearance despite local failures of the plating.

Jan 1, 1935

By C. A. Rowland, R. C. Nealey

Until recently there were very few dry grinding rod mills used. The early installations performed as anticipated. In the last seven years a number of dry grinding rod mills have been made, most of which have been used for grinding iron ore. The recent installations have not operated as anticipated in accordance with practices learned from wet rod milling. The paper does not establish any definite rules for the application of dry grinding rod mills but does attempt to establish some guidelines that can be followed. The paper reviews a number of flowsheets describing the problems encountered and some of the solutions found. Completely dried compared to partially dried feed is reviewed. Problems encountered in scaling up from pilot plant to production-size mills are described. Rod sizing, effect of rod length, rod bending, rod breakage, and rod tangling are discussed. While wet grinding rod mills have been used for 50 years to grind ores and make specifications sands, dry grinding rod mills have only recently come into significant use in the mineral industries. Among the materials now being processed in dry grinding rod mills are (1) lead sinter, lead slag, and limestone; (2) iron ore; (3) silica; (4) cement clinker; and (5) coke. Because the application of wet grinding rod mills was well understood at the time that dry grinding rod mills were introduced, the criteria and guidelines for selecting the former were carried over to the selection of the latter - in some cases with the anticipated results, but in many cases not. It is the purpose of this paper to describe, through a series of flowsheets, some of the recent problems encountered with dry grinding rod mills, to point out the solutions to these problems, and to set forth some guidelines for the selection and application of these mills. When dry grinding rod mills came into use, the following facts about their wet grinding counterparts were already well established: (1) The performance of large-sized mills can be accurately predicted from the performance of rod mills 5 or 6 ft in diameter. Wet grinding mills as large as 14 ft in diameter (inside shell) by 20 ft long with rods 19.5 ft long can operate satisfactorily.

Jan 1, 1970

By P. K. Strangway, H. U. Ross

Briquets consisting of pure artificial magnetite, pure artificial hematite, and mixtures of the two were reduced by hydrogen in a loss-in-weight furnace at temperatures in the range 500° to 1000° . The rate of reduction of the pure hematite briquets increased continuously with increased temperature. In contrast, the pure nmgnetite briquets exhibited a pronounced rate ninimutn at about 700°C. Metallographic studies of partially reduced briquets rerlealed that, at this temperature, the he.matite samples reduced in a topo-chemical manner while the magnetite ones reduced uniformly throughout, and after partial reduction their cross sections contained a mixture of iron and unreacted wustite grains. No iron shells could be detected on the surfices of any of these uwstite grains. X-ray diffraction investigations indicated that these grains had a rzinimum lattice parameter when they had been formed at the rate rninimum temperature. Also, it was found that an activation energy of 41,000 cal per mole zoas required for reduction when only these wustite grains were present. Thus, it is suggested that the overall reduction rate of the rnagnetile su?nples at temperatures in the range influenced by the rate nzinirnum phenomenon was limited by the rate qf iron ion diffusion in the unreacted wustite grains. THE rate minimum phenomenon, which has often been observed when reducing iron oxides at a temperature of about 700°C, is one of the most interesting, yet unresolved, problems in the field of reduction kinetics. Basic principles of chemical kinetics and 'In some instance, a second rate minimum has been observed at about 900°C. Since most investigators are in agreement that this minimum is directly related to the transformation from a to y iron (which takes place at 911°C) and since it was not encountered during the present reduction tests, it will not be referred to in this vaver. fundamental laws of diffusion all agree that, as the temperature is increased, the rate of reduction should also increase. However, with certain ores, it has been found that their reduction rate actually decreases with an increase in temperature up to some value X where a minimum reduction rate is reached. With further temperature increases beyond X the rate becomes more rapid again. Temperature X is usually referred to as the "rate minimum temperature", while the overall type of behavior constitutes the "rate minimum phenomenon". This phenomenon has been reported by numerous investigators. They have found rate minima during the reduction of both artifiial' and natural374 magnetites and artificia15j6 and natural5" hematites. Rate minima have been observed when reducing high-purity material2 or low-grade ores,3'4 when studying particles in the micronsize range5 or relatively large agglomerates,g10 and during reduction with either hydrogen7 or carbon monoxide.11"2 Previously, this phenomenon has been attributed to many factors; these include sintering and recrystallization of the iron formed during reduction374 changes in microporosity of the ore upon redction,"" formation of dense iron shells around retained wustite grains,11716 and chem-isorption,17 to name only a few. However, most investigators who have reported a rate minimum merely speculated as to what seemed to influence it and they did not examine the fundamental causes. Consequently, the present experimental study was initiated in order to evaluate the basic factors which could be associated with this phenomenon. MATERIALS AND METHODS The experimental techniques, followed during this investigation, are similar to those which have been described previously.18 The chemically pure magnetic powder was prepared by partially reducing Fisher reagent-grade hematite with a gaseous mixture of carbon monoxide and carbon dioxide in a rotating-drum furnace. Three-quarter-inch diam cylindrical briquets which weighed about 12 g were formed from this magnetite powder and pure hematite powder. All of the briquets were sintered while they were slowly raised through the 1200°C hot zone of a vertical tube furnace. An argon stream was continually flushed through this furnace in order to prevent oxidation of the magnetite briquets, while in the case of the pure hematite briquets sintering was carried out in air. The sintered hematite briquets had a density of 5.06 g per cu cm while the density of the sintered magnetite briquets was 4.27 g per cu cm. The sintered briquets were reduced by purified hydrogen in a loss-in-weight furnace at temperatures in the range 500" to 1000°C. In all instances, the critical reducing gas velocity was exceeded and, in order to ensure that the results were reproducible, duplicate briquets of each type were reduced under each set of experimental conditions. A continuous record of the weight loss during reduction was obtained with the aid of a Statham transducer. The present experimental setup was capable of detecting a change in weight as small as 10 mg. Since a weight loss of over 2 g usually occurred during each reduction test, an accuracy of better than 0.5 pct of the total weight loss could be achieved. RESULTS AND DISCUSSION Reducibility Tests. In the first set of experiments, pure hematite and pure magnetite briquets were used.

Jan 1, 1969

By F. R. Hensel

THE alloy under discussion is a nickel-silicon bronze and is one of many age-hardening or precipitation-hardening FIG. I.-EFFECT OF QUENCHING AND AGING TEMPERATURES ON HARDNESS OF TWO TYPICAL NICKEL-SILICON BRONZES, NO. 2351 AND No. 2362. copper-base alloys. The hardening con-stituent is a nickel-silicon compound.1 The quenching temperature necessary to pro-duce hardening by a subsequent aging treatment is a function of the composition and extends from about 700'C. to ()50'C.; the aging temperature range extends from 400° to 600°C. The work carried out by the authors of this paper was confined to cast alloys having compositions within the ranges of 1.5 to 3.00 per cent Ni; 0.4 to 0.75 per cent Si; balance, copper.

Jan 1, 1940

By F. C. Farnham

THERE has been very little, if any, uniformity in the reference datum used for magnetometer surveys. It has been the practice to choose a base station at which the magnetic field is assumed to be normal and call the difference in field strength between a given station and this base the anomaly at the given station. For small-scale surveys latitude and longitude corrections may or may not be applied, but for large-scale surveys such corrections usually are applied. The United States Coast and Geodetic Survey maps showing lines of equal magnetic intensity are commonly used to determine the normal latitude and longitude variation by scaling values from the map for given points and computing the change per mile north and the change per mile east. This method has been quite satisfactory for isolated surveys, but when two surveys originating at different base stations were extended until they joined, difficulty arose, not only because of discrepancies between base-station values, but also because of differences between the latitude and longitude correction factors used for the separate surveys. It is the purpose of this paper to describe a reference datum for magnetometer surveys which, it is believed, will eliminate most of the difficulty arising from this source. In order to be satisfactory, a general datum for magnetometer sur-veys should be as close as possible to the observed field and still he based on some assumed distribution of magnet poles, in order that the "normal" value of any component of the field for any given point may be computed to as many significant figures as desired. Such a field has been com-puted, and a comparison of the intensity of its vertical component with the average vertical intensity for the United States as shown by the U.S.C.G.S. map of lines of equal intensity for 1935 is shown in Fig. 1.

Jan 1, 1939

By Martin G. Castro, Renato G. Bautista, Morton Smutz

During the last 25 years, multistage mixer-settlers have been used extensively for liquid ion exchange work. The mixer-settler has the advantage of being a compact device in which a large number of stages can be assembled in a relatively small area. They are particularly advantageous for the separation of rare earths. Most of the adjacent rare earths have separation factors which are near one, which means that a large number of stages are needed to produce a good separation. When two elements (m, n) are extracted simultaneously, the separation factor (ßm, n) is defined as: [ ] where Ym, Yn are the molarities of the respective elements in the organic phase and Xm, Xn, are their molarities in the aqueous phase at equilibrium. The industrial consumption of rare earth oxides rose to over 6800 tons in 1968, an increase of 33% over 1967.1 The increased production and demand for rare earth and lanthanum chlorides for petroleum cracking catalysts was one of the principal factors in this increased consumption of rare earths. In 1968 approximately 59% of the rare earth oxides produced were used for gasoline cracking catalysts in the form of rare earth and lanthanum chlorides. There was also an increase in the sales of lanthanum oxides for use in optical lenses and in fiber optics. The future looks good for the rare earth industry. The rare earths can be expected to be used in many new products which are being developed. Until the late 1950's tributyl phosphate (TBP) was the principal solvent used in rare earth liquid ion exchange. In 1957 Peppard3 reported the use of di (2-ethyl-hexyl) phosphoric acid (D2EHPA) as a solvent for separating the rare earths. Rare earth separation factors of up to 2.5 were obtained in nitrate and chloride systems while the maximum separation factor using the

Jan 1, 1972

By Roy A. Lindgren

SINCE the year 1643, when the first blast furnace in America for treating iron ore was built at Saugus, Mass., out of mica schist quarried in the neighboring district, the procurement of a suitable refractory for furnace lining has been a problem of concern to the operators of furnaces. The stacks built of mica schist continued to smelt iron ore until about 1836, when, according to F. H. Norton, the first firebrick were produced', at Queens Run, Pa. Other writers speak of brick having been molded and burned in Massachusetts about the year 1834. In 1841, Andrew Russell began to produce medium refractory plastic clay brick near East Liverpool, Ohio, that were used for lining blast furnaces'. The well-known Kentucky clay-producing district was not opened up until the year 1871, but since then it has produced a large percentage of the linings for iron blast furnaces. While some strides have been made by the refractories industry during the 100 years that have passed since the first firebrick were produced, it has been only during the last two decades that any real progress has been made towards bettering the product, even though the method of production had improved. Perhaps the fault lies with the user of the brick rather than with the producer, for not sooner demanding a supe¬rior product. During the past 15 years the tonnage produced per lining has increased from 500,000 gross tons to 1,000,000 gross tons, and now some furnaces are producing 1,600,000 gross tons and better on a single lining. It is true that enlarged capacity of furnaces and improved practice have accounted for some of this increase in tonnage, nevertheless better quality in firebrick must be given credit for its share. However, we are not yet ready to say that we have reached a maximum life of furnace lining. We believe that a better product can be produced and that the refractories industries of America will, through their extensive research depart-

Jan 1, 1937

By V. Annamalai, J. B. Hiskey

Abstract-The kinetics of copper cementation on a rotating disc of pure aluminum were studied as a function of chloride ion concentration, rotational speed of the aluminum disc, hydrogen ion concentration, copper ion concentration, and temperature. The optimum experimental conditions for all the parameters were determined. Chloride ion was found to destroy the naturally existing aluminum oxide film, and enabled the cementation reaction to proceed. A two-step rate, observed in most of the experiments, was attributed to the attainment of a roughened surface. An Evans diagram for the copper-pure aluminum system was constructed and utilized in predicting the controlling mechanism of the system. At low temperatures (<33°C), the rate with an experimental activation energy of 20.36 kcal mol-1 was controlled by the presence of an oxide layer. While at higher temperatures presence of an oxide layer. While at higher temperatures (>33° mol-1 was controlled by ionic diffusion. The deposits obtained were pure copper and not a copper-aluminum alloy. The amount of aluminum dissolved was determined for each experiment, and the maximum excess consumption was found to be only 40%. The structural characteristics of the copper deposits were studied using a scanning electron microscope. These observations indicated that the deposits were either bulbous, massive crystalline, or dendrites of different degrees of fineness depending on the experimental variables. The reaction rates seemed to depend on the deposit structures in a predictable way when a particular experimental parameter was varied.

Jan 6, 1978

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Jan 1, 1925

By Kenneth A. Gutschick, Robert S. Boynton

Lime has become a general and loosely used term to denote almost any kind of calcareous material or finely divided form of limestone or dolomite, as well as burned forms of lime. However, according to Webster, lime is only calcined limestone (known as quicklime, unslaked lime, or calcium oxide). It also embraces the secondary product of quicklime, i.e., hydrated lime or slaked lime (calcium hydroxide). There are two basic types of limestone used for lime manufacture-high calcium and high magnesium (dolomitic). The dolomitic quicklimes and hydrated limes correspond to their "high calcium" counterparts. The only difference is that the dolomitic types are a combination of the elements calcium and magnesium, in varying percents, whereas high calicum limes contain less than 5% MgO down to 1/2%. In either case a pure form of limestone of at least 97% combined carbonate content is needed to make salable lime, barring a few exceptions. The calcination of both types is diagrammed chemically as follows: [ ] In the foregoing reversible reactions the limestone is burned (calcined) in lime kilns, with the carbon dioxide content of the stone expelled as a gas. The weight loss during calcination of pure high calcium limestone is 44%; with equimolecular dolomitic limestone, the weight loss is 48%. Since lime has a strong affinity for CO_, particularly if moisture is present, it will readily revert to its original carbonate form. This occurs when quicklime "air-slakes" and adsorbs CO.. from the air. Thus, quicklime is perishable and should be stored in dry, water-tight areas. Depending upon the type of kiln used and the physical structure of the stone, the size of the quicklime may range from sandlike granules to 8 in. lumps. However, pebble sizes ranging from 1/4 in. to 2 in. are the most common. Screened quicklime fines are also com¬pressed (pelletized) into 1 in. pellets. For some purposes ground or pulverized forms of quicklime are used, ranging from No. 10 mesh to dust. Generally a pebble of quicklime is about the same size as the pebble of limestone before calcination (perhaps shrinking slightly). A more stable form of lime is hydrated lime. This is obtained by adding water to quicklime, slaking it into a dry, fine, subsieve size white powder. Because its affinity for moisture has been satisfied, it is not vulnerable to further moisture like quicklime. However, it has the same strong affinity for CO_. The following reversible chemical reactions illustrate how hydrated lime when heated (or dehydrated) can revert to the original oxide form:

Jan 1, 1975

By H. E. Swanson

A description of the field apparatus has been published by D. G. Brubaker. Data from laboratory model studies of the in-line and broadside methods of operation are detailed. The conductor models include a single sheet conductor at several strike and dip angles and a schistose-type conductor. In addition, data on the effects of the strike length and of the depth of the conductor are presented for the broadside method of operation. All of the data show that the inclintion reverses direction over the top edge of dipping single sheet conductors. Differences between anomalies over conductors dipping between 90" and 30" are subtle,but flat-lying sheets can be readily distinguished from steeply dipping sheets. Schistose conductors are easily distinguished from single sheet conductors, and a procedure is given for determing the direction of schistosity. The depth of penetration under ideal conditions is 0.7 of the coil separation. In 1950, The New Jersey Zinc Co. (of Pa.) developed a lightweight apparatus for electromagnetic prospecting. A description of this apparatus has been published by D. G. Brubaker.1 It consists of a small, battery-powered vertical source coil and a direction-finding receiving coil. Source and receiver are kept a fixed distance apart and move as a unit along the survey lines. At each receiver station the inclination of the magnetic field created by the source is measured. Two methods of operation are used, in each of which measurements are taken with a constant source-to-receiver distance and with the plane of the source oriented to pass through the receiver station. In one procedure, called the in-line method, source and receiver travel in tandem along the same line. In the other procedure, called the broadside method, source and receiver move along adjacent lines, with the source-to-receiver direction oriented parallel to the assumed strike direction. The data recorded in either case are the inclinations, measured in degrees from the horizontal, of the magnetic field created by the source. These data are plotted at the receiver stations on a map. Scale-model studies of both methods of operation have been conducted in the laboratory and are reported in this paper. APPARATUS Small source and receiving coils wound with many turns of fine copper wire were used to produce and measure an alternating magnetic field in the same manner as the field apparatus. The model source coil was mounted in a vertical position and oriented so that the receiving coil was in the plane of the source. The receiving coil was mounted in a frame so it could be rotated to obtain the inclination measurements. The source coil was excited by a 100-kc signal generator. A tuned amplifier and a voltmeter were used to detect the null position of the receiving coil. The coils used in the field apparatus are usually spaced about 400 ft. The coils used in the model work were spaced from 4 in. to 4 ft apart. Thus the scale of the model work ranges between 1/1200 to 1/100 natural size. The model conductors were aluminum-coated building paper supported on wooden frames. The conductivity-thickness product was measured using a four-electrode system and was found to be about 300 ohms-&apos;. CONDUCTIVITY CONSIDERATIONS The electromagnetic response from a thin sheet conductor is a function of the product of the conductivity of the sheet times the thickness of the sheet times the frequency of the alternating electromagnetic field times some linear dimension of the system such as the coil separation which indicates scale. The secondary electromagnetic field produced by the conductor increases in amplitude from zero at zero conductivity to a maximum value for infinite conductivity, while the phase of the secondary field with respect to the primary field shifts from 90" at zero conductivity to 0" at infinite conductivity. Several authors2-&apos; have shown that most of both the increase in amplitude and the change in phase of the secondary field takes place over a three-decade range of the conductivity-thickness-frequency product. Only small changes in the amplitude and

Jan 1, 1961

By K. R. Janowski, H. Conrad

As part of a program to establish the effect of crystal imperfections on laser output, a detailed study was made of the dislocation structure of ruby crystals obtained from varioius sources. Using KHSO4 as an etchant, a detailed mapping of the dislocation structure on the (0001) and the (1150) planes was carried out. A less extensive study was made of the rhombohedral planes. The average dislocation density on the basal plane was 1.5 to 3 x 106 cm&apos;2 and on the (1120) planes -5 x 10&apos; cm-2. Howe7!er, considerable variation existed between areas on a given plane. The subboundaries in the basal plane tended to lie along [1100] type directions while those on the (1120) plane tended to lie: a) parallel to the basal plane, b) along truces of the (1101) plane, and c) along normals to the traces of the (0001) planes. The orientations suggest that mang of the dislocations lie on the (0001), (1701), and (1150) planes A Laue back-rellection nnalysis Of c-axis wander in one crystal revealed a gradual increase in misorientation from me end of the crystal to the other, with a maximum variation of 1°35&apos;. Slight etching of a mechanically polished crystal revealed mumerous polishing scratches, suggesting the existence of a plastic flow layer. It is recognized that chromium concentration, lineage, and other crystalline imperfections play a role in the efficiency of ruby lasers.&apos; Consequently, it is desirable to establish in detail the imperfection structure of ruby crystals to be used as lasers. Some exploratory work along this line had already been done by Barnes&apos; and Dils and Martin.&apos;" How- referred to as the morphological system, as described by Kronberg.4 The Miller indices given in the present paper will be based on this morphological system. In this system, the three common growth planes are the (0001), the {1120}, and the {1101}.5 Since these planes are of high stability, it is expected that they are most suitable for study by the etch-pit technique. In the present investigation, a detailed mapping of the dislocation structure was carried out for the (0001) and the (1130) planes. A less extensive study was made of the etch-pit structure on rhombohedra1 planes. An analysis of dislocations in sapphire (the ruby host) was first made by kronberg,4 who postulated that slip on the (0001) plane occurred by motion of dislocations with Burgers vector 1/3 <1150> and that these dislocations might dissociate into quarter partials. An etch-pit technique using boiling phosphoric acid was developed by Scheuplein and Gibbs6, 7 for identifying dislocations intersecting the (0001) planes and the planes near (2021) in sapphire. Al-ford and Stephens8 were able to reveal dislocations on the (0001) and {1120} planes in sapphire and ruby by etching in fused potassium bisulfate (KHS04) at 675°C. The termini of basal dislocations [(000l), <1120>] were revealed by Palmour and Kriegel9 on {1120} and {1010} by thermally etching sapphire. Bond and Harvey10 were able to decorate dislocations in sapphire by adding 1 pet by weight of ZrO2 to the Al2O3 powder used in the growth of the crystals by the Verneuil technique and then heating the grown crystals for 4 hr at 1500°C. They found, using optical transmission microscopy, that many dislocations had straight segments which were parallel to one of the six [1120] directions in the basal plane. Single dislocation lines curved smoothly from one alignment to the other, turning through angles of 60 to 120 deg. Also observed were isolated hexagonal loops, approximately 10 µ in diameter, the sides of which lay along the same directions as did the longer dislocations. MATERIAL AND ETCHING PROCEDURE The ruby specimens used in this study (all containing approximately 0.04 pet Cr) had been obtained from four different sources: Linde Co., Meller Co., Valpey Crystal Corp., and those grown at Aerospace Corp.* For the initial phase of this study, several specimens were cut from the slab shown schematically in Fig. 1 which had previously been cut from a disk boule obtained from the Linde

Jan 1, 1964

By R. A. Meussner, C. T. Fujii

Chromium solution in wustite depresses the oxygen activity in a nonideal manner and expands the lattice slightly. Gravimetric measurements of the equilibrium compositions of wustite containing 0.00 to 1.38 wt pct Cr define the oxygen potential (CO2-CO atm) and the limits of the phase field at 1000°C. These data extrapolate to a maximum solubility of 2 wt pct Cr. The lattice parameter data, room-temperature measurements of quenched samples, differ significantly from those in the literature. Both pure and chromium wus-tites contract at uniform and equal rates as the oxygen content (metal deficit) increases. The a. US metal deficit relationships are straight lines showing none of the curvature previously reported. At constant metal deficit, the a. expands by 0.002? on alloying with 0.4 to 0.5 wt pct Cr but is insensitive to further chromium additions: a small expansion rather than a marked contraction. These results require a modification of the accepted alloying mechanism. IN the high-temperature oxidation of Fe-Cr alloys in Ha-H,and CO2-CO atmospheres the vapor transfer of oxygen from the continuous wustite outer scale to the porous inner scale/alloy interface sustains the high oxidation rates.1"3 The driving force for this transport is the difference in the oxygen activity of the outer wustite and that at the inner scale/alloy interface. In pure CO2, as well as in CO2-CO atmospheres, carburization of the alloy accompanies this oxidation. Current efforts to evaluate the parameters controlling this carburization have shown that the process is not simple; i.e., the carbon concentration in the alloy initially increases rapidly, reaches a maximum, and then decreases slowly as the oxidation time is extended. This is the same pattern reported by McCoy for the more complex oxidation-carburization of stainless steels and an Fe-Cr alloy in CO2 at lower temperatures.4 These changes in the carburization process occur while the overall oxidation rate and the lattice parameters of the scale layers remain constant. If this invariance of the lattice parameter is assumed to indicate a constant scale composition and thus a constant oxygen potential, the carburization results are not easily explained. Electron microprobe traces across the thickness of these outer scales, however, have revealed distinct chromium gradients penetrating 10 to 50 µ from the inner surface where the chromium concentrations were estimated at a few tenths percent. The variation of the chromium cmcentration on the inner surface of the scale, and the resulting change of the oxygen activity, during the oxidation process is considered to be important in defining the carburization process. Thus, to gain an understanding of the complex oxidation-carburization proc- ess, it was necessary to measure the properties of wustite containing known amounts of chromium (chromium wustite). Although the general features of the Fe-Cr-0 equilibrium diagram between 900° and 1300°C have been fairly well established by Richards and white: Wood-house and White,6 Seybolt,7 and Katsura and Muan,8 there have been no detailed studies of the limits of the chromium wustite phase field or the properties of these alloyed wustites. The recent results of a limited study of chromium wustite by Levin and wagner9 indicate that more than 0.67 wt pct Cr is soluble in wustite above 850°C and that this alloying is accompanied by a substantial lattice contraction. This change in the lattice parameter was not evident in the X-ray diffraction patterns of wustite layers from the oxidation experiments;1,3 the lines were sharp even though a chromium gradient existed in the scale. The present paper describes gravimetric equilibrium experiments which delineate the boundaries of the chromium wustite single-phase field at 1000°C and define the changes in the oxygen activity and the lattice parameter of wustite as functions of chromium and oxygen contents. The lattice parameter data of chromium wustite obtained from these equilibrium and special quenching experiments differ considerably from those reported.9 Those for pure wustite show significant differences from the widely accepted data of Jette and Foote.10 Since the causes of these differences are not easily assignable, and since the literature contains many sets of data on wustite which if not in conflict are at best not in harmony, the experimental procedures are discussed in the present paper. EXPERIMENTAL Oxide Preparation. The specimens used in these studies were porous pellets compacted from high-purity (Fe,Cr)2O3 or Fe2O3 powders. These powders were prepared from electrolytic chromium (99.8 pct purity) and electrolytic iron purified by consumable-electrode, vacuum arc-melting processing (299.9 pct purity).11 The oxides were produced by standard analytical procedures: solution of the metals in acids (HC1, then HNO3 added), coprecipitation of the hydroxides (NHaOH), and filtering and washing these precipitates. Initial dehydration of the coprecipitated hydroxides was done in a porcelain evaporating dish heated by a Meker-type burner, final dehydration by firing in a recrystallized alumina crucible at 1000" to 1050°C for 20 hr in an electrically heated furnace. Spectrographic analysis detected the following impurities: silicon, 0.01 pct or less; aluminum, 0.001 pct to trace; manganese and copper, 0.0001 pct to none. Each of these levels is at least one order of magnitude lower than in commercially available CP grade ferric and chromic oxides. Fluorescence analy-

Jan 1, 1969

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons introduced by members) Member, aged 32. Fifteen years&apos; experience in the mining, quarrying, and manufacture of gypsum and gypsum products, as purchasing agent, designer, operating manager and managing executive. Thoroughly familiar with the business in all branches, especially the organization of efficient producing forces and machinery and the value of cost and their application to manufacturing. Desires responsible position with a financially reliable concern in the same or similar business in which an interest may be obtained at a later date. Now doing consulting work, but free for appointments after April 1. New York or Philadelphia. No. 345. Member, technical graduate. Unmarried. Aged 32. Seven and one half years&apos; experience in mining, timbering, and mine sampling. At present, and for past two and a half years, mine sampler for well known and highly successful mining company. Desire change. Go anywhere. Prefer assisting Mining Engineer. No. 346. Member, 40 years old. 18 years&apos; experience in examinations and operation in the United States, Mexico and South America. Especially successful with poor class of labor. Open for engagement in any part of the world, in technical, administrative or consulting work. No. 347. Member, technical graduate. Sixteen years&apos; experience as engineer, superintendent, and manager in metal mines. Exploratory . work, developing and equipping new properties. Operation of large mines, and various systems of mining. Low costs. Specialized in iron and copper mining, but have had some experience in zinc, lead and pyrite mining. No; 348.

Jan 4, 1917

By P. van Meurs, C. van der Poel

From observations of the linear displacement of oil by water from a porous medium as visualized in transparent models, new insight into the mechanism of vircous fingering, as occurring in the case of unfavorable oil-water viscosity ratios, has been obtained. A mathematical description of the observed phenomena leads to formulas. expressing both oil production and pressure drop across the formation as a func-tion of cumulative water injection with the oil-water viscosity ratio as a parameter. Theoretical results are compared with those obtained from scaled laboratory experiments in sand-filled rubes in which water was injected at the one end and production was obtained from the other end. In the second part of the paper the same principles are used to calculate a more realistic production scheme in which production of oil and water is obtained from wells regularly distributed over the field. Results show satisfactory agreement with production data of an actual oil field. Application of the theory is extremely simple as only a limited number of readily available parameters have to be known and results are presented in an analytical form. INTRODUCTION When oil is displaced from a porous medium by water having a viscosity lower than the oil. the oil-water interface is essentially unstable and has the tendency to break up into what are called "fingers" or "streamers". This Viscous fingering has been noted before by Engelberts and Klinkenberg&apos; and has recently been demonstrated in transparent models by van Meurs-. It is highly questionable whether it is at all possible to perform relative permeability measurements In the laboratory reproducing the same distribution of water and oil in the cores as Present in practice. Therefore, we based our descriptive theory on an idealized model which was "tailored" to the fingering mechanism as observed in transparent models under scaled conditions. This approach was the more attractive, as a highly simplified model had already yielded an adequate description of the process. Though we are aware our theoretical approach can be extended in several ways, it is given here in its simplest form to explain the fundamental principles and to avoid complicated formulas. SIMPLEST CASE OF A LINEAR DRIVE Observations and Theoretical Model In Fig. 1 a series of photographs shows how water penetrates into a formation saturated with an oil which is considerably more viscous than the water. The technique for obtaining these photographs, using transparent models, has already been described?. The fingers can be clearly observed and are seen to be mainly parallel to the flow direction. Note that the contours of these fingers are not smooth but show protrusions of water enclosing oil pockets.

By Roy Lindgren

Since the year 1643, when the first blast furnace in America for treating iron ore was built at Saugus, Mass., out of mica schist quarried in the neighboring district, the procurement of a suitable refractory for furnace lining has been a problem of concern to the operators of furnaces. The stacks built of mica schist continued to smelt iron ore until about 1836, when, according to F. H. Norton, the first firebrick were produced1, at Queens Run, Pa. Other writers speak of brick having been molded and burned in Massachusetts about the year 1834. In 1841, Andrew Russell began to produce medium refractory plastic clay brick near East Liverpool, Ohio, that were used for lining blast furnaces1. The well-known Kentucky clay-producing district was not opened up until the year 1871, but since then it has produced a large percentage of the linings for iron blast furnaces. While some strides have been made by the refractories industry during the 100 years that have passed since the first firebrick were produced, it has been only during the last two decades that any real progress has been made towards bettering the product, even though the method of production had improved. Perhaps the fault lies with the user of the brick rather than with the producer, for not sooner demanding a supe¬rior product. During the past 15 years the tonnage produced per lining has increased from 500,000 gross tons to 1,000,000 gross tons, and now some furnaces are producing 1,600,000 gross tons and better on a single lining. It is true that enlarged capacity of furnaces and improved practice have accounted for some of this increase in tonnage, nevertheless better quality in firebrick must be given credit for its share. However, we are not yet ready to say that we have reached a maximum life of furnace lining. We believe that a better product can be produced and that the refractories industries of America will, through their extensive research depart-

Jan 1, 1936

By Leon Merker

Flame fusion growth of strontium titanate crystals was undertaken to obtain large transparent crystals on which physical data could be gathered. The fact that strontium titanate is a cubic crystal and probably has a high index of refraction indicated that it might be useful in optics. No large boules* of strontium titanate had been synthesized previously, and none have been found as natural occurrences. The flame fusion process, invented by Verneuil in 1904, was first applied to grow large crystals of ruby and has since been used extensively to grow crystals of corundum and spinel on a commercial scale. Only during the past nine years has this technique been applied to the growth of large single crystals of other high melting compounds such as rutile, mullite and ferrites.

Jul 1, 1955

Extremes in temperature and weather, along with the highly corrosive nature of acid leach solutions used at open-pit copper mines, necessitates the use of pipeline systems that are both corrosion resistant and impervious to the environment.

Jan 6, 1976

By P. D. Wilson

The genesis of some orebodies has been explained by the mingling and chemical interaction of water solutions of different compositions and the consequent precipitation of the mineral load of one or both solutions at their confluence. While this process can be proved in the laboratory, rarely can a positive demonstration be found in nature, for the conditions at the time an ore deposit is being worked are usually entircly differcnt from those that existed when the orebody was formed. For this reason, the present deposition of copper carbonate in the southern end of the Briggs mine of the Calumet and Arizona Mining Co. near Bisbce, Ariz., is of great interest. The developed orebodies in the Calumet and Arizona mines are roughly confined to two broad parallel zones separated at the north by a wedge of Copper Queen ground and at the south by an undeveloped block of Calumet and Arizona ground about 2000 ft. (610 m.) wide. The more westerly zone includes the Irish Mag, Oliver, P. and D., and Cole mines, none of which has been operated since 1919. Much of the water from this division is carried from the 1400-ft., or bottom, level at the southern end of the Cole mine in a long water drift through the undeveloped block to the 1300-ft. level of the Briggs mine at the southern end of the more easterly zone. Most of this water comes originally from outlying areas in the P. and D. mine, passing through parts of the lower. levels of the Cole mine, which are heavily mineralized with iron and copper sulfides. This sulfide zone has been opened up and slowly oxidizing for years. Originally a pure, potable water, the analyses of which are shown under A- in Tables 1 and 2, on its way through the mineralized area in the Cole mine it gathers copper, calcium, magnesium, iron, and aluminum sulfates, a load of dissolved mineral salts almost four times as great as that which it formerly carried; analyses B

Jan 1, 1923

By P. D. Wilson

The genesis of some orebodies has been explained by the mingling and chemical interaction of water solutions of different compositions and the consequent precipitation of the mineral load of one or both solutions at their confluence. While this process can be proved in the laboratory, rarely can a positive demonstration be found in nature, for the conditions at the time an ore deposit is being worked are usually entircly differcnt from those that existed when the orebody was formed. For this reason, the present deposition of copper carbonate in the southern end of the Briggs mine of the Calumet and Arizona Mining Co. near Bisbce, Ariz., is of great interest. The developed orebodies in the Calumet and Arizona mines are roughly confined to two broad parallel zones separated at the north by a wedge of Copper Queen ground and at the south by an undeveloped block of Calumet and Arizona ground about 2000 ft. (610 m.) wide. The more westerly zone includes the Irish Mag, Oliver, P. and D., and Cole mines, none of which has been operated since 1919. Much of the water from this division is carried from the 1400-ft., or bottom, level at the southern end of the Cole mine in a long water drift through the undeveloped block to the 1300-ft. level of the Briggs mine at the southern end of the more easterly zone. Most of this water comes originally from outlying areas in the P. and D. mine, passing through parts of the lower. levels of the Cole mine, which are heavily mineralized with iron and copper sulfides. This sulfide zone has been opened up and slowly oxidizing for years. Originally a pure, potable water, the analyses of which are shown under A- in Tables 1 and 2, on its way through the mineralized area in the Cole mine it gathers copper, calcium, magnesium, iron, and aluminum sulfates, a load of dissolved mineral salts almost four times as great as that which it formerly carried; analyses B

Jan 1, 1923