Search Documents

By Arthur Wade

The most important development in connection with the search for oil in Australia during 1940 was on the legislative side. Prior to 1939 legislative enactments in the various states with regard to prospecting for petroleum were such as to prohibit large-scale effort and to bar the major oil companies from operating. The preparation of the Wade Model Bill for the Commonwealth Government in 1939 and the acceptance of the Bill in principle by all States, has gone far to remove these obstacles. Amending Acts, largely based on the Alodel Bill, have now been passed by all States with the exception of New South Wales and Tasmania. In New South Wales an Amending Bill is being drafted, while the Govern-ment of Tasmania is preparing to introduce an Amending Bill provided there is any demand for it. Noteworthy developments followed immediately on the passing of these Amending Acts. The Shell Company (1)1 and the Superior Oil Company of California (2) too$ up large areas in Queensland and com-menced work during 1940. Caltex acquired a big tract of country in the Kimberley district of Western Australia (3) and has made application for an area in the Territory of Papua. An Australian company, The Phoenix Oil Extraction Co., has obtained a permit to prospect over an extensive area bordering the south coast of Western Australia (4). Other developments are foreshadowed. The Shell Company is making an intensive examination of its area in Queensland (I). Geological, geophysical and aerial surveys are in progress. Caltex has three geological parties in the held in Western Australia while Superior Oil Company of California has done some geological work in the interior of Queensland (2). None of these concerns envisages a drilling campaign in the near future. Much exploratory work is necessary. Further work has been done by the companies that were already operating in 1939, though an important Australian company - Oil Search—has abandoned the work of prospecting in Australia. This company has subleased its holdings, at North West Cape in Western Australia (;), to Caltex on a royalty basis. The company also maintains its holdings in, and connection with, the Australasian Petroleum Company Proprietary Limited, which is a combination of Standard Oil (New Jersey), Anglo-Iranian and Oil Search. This company is operating in the territories of Papua and New Guinea. Drilling Victoria.—Although a good deal of drilling was done in the LaBes Entrance Tertiary Basin of Victoria (7), no appreciable production of oil was obtained. In order to test the extent and possibilities of the oil-bearing sand in this region, the Commonwealth and State Governments jointly put down six scout boreholes to depths of between 1200 and I500 it. The results were nega tive. A deeper exploratory horehole put down by the Governments at Seacornbe, or 26

Jan 1, 1941

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On' (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By Walter R. Hibbard

ALTHOUGH extensive pole figure studies have been reported by Brick1 and others2,3 showing preferred orientations in rolled and annealed 70-30 alpha brass, and by Hermann and Sachs4 in 70-30 alpha brass cups, comparatively little information has been published on textures in drawn 70-30 alpha brass tubes. Crampton5 reported a (110), [III] texture in drawn brass tubes based on selected individual X ray photograms following the method of Norton and Hiller,6 but the extent and ease of developing this texture and the annealing texture have not been reported and pole figures of these structures were not found in the literature. The present investigation was further stimulated by the possible influence of textures on the mercury stress-corrosion cracking of brass tubes. According to Snoek,7 iron-nickel alloy sheet with a strong cubic texture was more resistant to corrosion than a randomly oriented sheet. Crampon5 reported that single crystals of 70-30 alpha brass drawn into tubes did not crack after 24 hr s in a standard mercurous nitrate test. Edmunds 12 found that single crystals of 70-30 alpha brass did not crack in a standard mercurous nitrate test but did crack when subjected to ammoniacal vapor. Wassermann reported a similar failure of a single crystal of brass in ammonia. A tube with a strong texture might approach single crystal behavior and tend to resist mercury cracking if the differences in orientation between adjacent grains were sufficiently slight. PREPARATION OF SPECIMENS Crampton's5 tubes drawn 28 pct reduction in area developed a slight degree of preferred orientation while in tubes drawn 48 pct it was more marked. Based on this information, in order to obtain specimens with markedly different degrees of preferred orientation but with the same final draw, two sets of 70-30 alpha brass tubes were made following the mill schedules shown in Table I. The final draw involved a large reduction in outside diameter, and a ratio of reduction in area to reduction in outside diameter similar to that reported by Crampton,5,10 to cause mercury cracking of brass tubes. The brass contained 71.31 pct copper, 0.01 pct lead, 0.01 pct iron and zinc by difference. DETERMINATION OF POLE FIGURES A 6-in. length from each tube was rotated rapidly by an electric stirrer in 25 pct nitric acid until the gauge was reduced to 0.003-0.005 in. The tubes were carefully cut longitudinally and the walls gently flattened. They were further etched in 25 pct nitric acid to 0.0015 in. thickness and (III) pole figures were determined using copper K-alpha radiation in the manner described by Brick.1 The three general types of pole figures

Jan 1, 1947

By K. S. Krishnam, A. G. Guy

VOLUME 6 of the Transactions of the Baikov Institute of Metallurgy, Moscow, is devoted to the subject of "the experimental and theoretical study of the physical characteristics of dilute solid solutions,11 and summarizes work that has been done largely in the years 1958-9. In view of the attention that Soviet investigators have given to this subject, it seemed useful to carry out an independent experimental investigation in an effort to judge the scientific basis and practical significance of this field of study. The present note is a preliminary report of this investigation. The first property studied by the Soviet group was X-ray spectra of solid-solution alloys.' An anomaly observed at low concentrations (--0.05 at. pct) of the alloying element led to the theory that each impurity atom in a transition-metal base alloy is the center of a "block" of affected atoms with a radius of about 10"7 cm. This theory was further developed in a later paper.2 By this time similar anomalies had been found in diffusion coefficients3 and in thermal expansion.4 The entire development up to 1959 was reviewed in a comprehensive paper, and further theoretical ideas were also published at this time. Finally, the volume mentioned above was published in 19607 giving new experimental data on anomalies in optical properties and Hall constant, and a new theoretical treatment also appeared.' Although western scientists have done both theoretical work, summarized recently by Friedel,9 and considerable experimental work on dilute solid solutions, there is little support for the marked anomalies reported in the Soviet papers. Of the physical properties studied by the Soviet investigators, thermal expansion was chosen for the present investigation because of the relative ease of obtaining the necessary precision of measurement. A Leitz dilatometer was used for this purpose, and a special technique involving interchanging of the two specimens being compared permitted a precision of 0.05 x X in average thermal expansion over the range 27° to 600°C. The specimens were heated and cooled in vacuum or in a protective atmosphere at a rate of 50 °C per hr. Ferrovac E iron was used as the comparison specimen and in every case had a coefficient of expansion higher than (or equal to) that of the alloy. Since unintentional impurities are an important factor in studies of dilute alloys, special attention was given to the purity of the iron and the alloys used. The alloys were the same of those used by Abrahamson10 and by Michalak," and their preparation is described in these references. Chemical analyses of the Ferrovac E iron and of typical Fe-W and Fe-Mo alloys are given in Table I. The purity of the iron and alloys is undoubtedly higher than in the

Jan 1, 1962

By O. Schaaber

THE existence of isothermal martensite without the presence of the athermal form thus far has been established only in a 6 pct Mn, 0.6 pct C steel' and a 23 pct Ni, 3.5 pct Mn-Fe alloy.' " Theoretical calculations, moreover, predict this phenomenon to be observable only within narrow composition limits.' In all other steels, it is generally considered that martensite inevitably forms upon cooling below the M, temperature. Even the possibility of isothermal martensite transformation has been questioned for the common steels, although Averbach and Cohen" showed that martensite formation does not stop when the cooling is arrested and that as much as 5 pct retained austenite may transform isothermally to martensite at room temperature. This note presents some observations made on a number of low alloy steels by dilatometric and inductive measurement of isothermal austenite transformation at various temperatures close to the Ms. In the dilatometer studies reported previously, austenitized specimens were immersed in a salt bath maintained at proper temperature and immediately inserted into the dilatometera8' already at thermal equilibrium with the bath. The dilation was read on a dial gage divided into 0.001 mm intervals. Induction measurements were made with a separate excitation source and a test coil, sealed together in a salt-tight case. The increase in the induced voltage caused by the transformation products was recorded at a paper speed of 4 mm per sec. The steels studied are listed in Table I. Specimens prepared from steels 1 and 2 were in the form of hollow cylinders 60 mm long, 10 mm ID, 14 to 16 mm OD, while specimens of steels 3 and 4 were in the form of notched-bar impact test specimens, DVM type. Specimens of steel 5 were in the form of strip, 7x0.3 mm. In the bainitic range, curves representing the amount of transformation vs log time have a single step appearance, as is shown by the curve of isothermal transformation at 320°C in Fig. 1.Y ransformation proceeds in accordance with the Austin-Rickett equationo P -----------= k t" 100-P where P is the percentage of transformed austenite, t is time, and k and n are constants. Near the M, temperature, however, complex curves having two steps are observed. As shown in a previous papera for steel 3 isothermally transformed at 245"C, such complex curves can be separated into two separate transformations or reactions each following the Austin-Rickett equation. The values of n in the equation for a number of such two-step transformations are listed in Table 11 for steel 1. At transformation temperatures of 250" and 235"C, evidence of a second transformation is very

Jan 1, 1956

By Arthur Wade

The most important development in connection with the search for oil in Australia during 1940 was on the legislative side. Prior to 1939 legislative enactments in the various states with regard to prospecting for petroleum were such as to prohibit large-scale effort and to bar the major oil companies from operating. The preparation of the Wade Model Bill for the Commonwealth Government in 1939 and the acceptance of the Bill in principle by all States, has gone far to remove these obstacles. Amending Acts, largely based on the Alodel Bill, have now been passed by all States with the exception of New South Wales and Tasmania. In New South Wales an Amending Bill is being drafted, while the Govern-ment of Tasmania is preparing to introduce an Amending Bill provided there is any demand for it. Noteworthy developments followed immediately on the passing of these Amending Acts. The Shell Company (1)1 and the Superior Oil Company of California (2) too$ up large areas in Queensland and com-menced work during 1940. Caltex acquired a big tract of country in the Kimberley district of Western Australia (3) and has made application for an area in the Territory of Papua. An Australian company, The Phoenix Oil Extraction Co., has obtained a permit to prospect over an extensive area bordering the south coast of Western Australia (4). Other developments are foreshadowed. The Shell Company is making an intensive examination of its area in Queensland (I). Geological, geophysical and aerial surveys are in progress. Caltex has three geological parties in the held in Western Australia while Superior Oil Company of California has done some geological work in the interior of Queensland (2). None of these concerns envisages a drilling campaign in the near future. Much exploratory work is necessary. Further work has been done by the companies that were already operating in 1939, though an important Australian company - Oil Search—has abandoned the work of prospecting in Australia. This company has subleased its holdings, at North West Cape in Western Australia (;), to Caltex on a royalty basis. The company also maintains its holdings in, and connection with, the Australasian Petroleum Company Proprietary Limited, which is a combination of Standard Oil (New Jersey), Anglo-Iranian and Oil Search. This company is operating in the territories of Papua and New Guinea. Drilling Victoria.—Although a good deal of drilling was done in the LaBes Entrance Tertiary Basin of Victoria (7), no appreciable production of oil was obtained. In order to test the extent and possibilities of the oil-bearing sand in this region, the Commonwealth and State Governments jointly put down six scout boreholes to depths of between 1200 and I500 it. The results were nega tive. A deeper exploratory horehole put down by the Governments at Seacornbe, or 26

Jan 1, 1941

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On' (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By Herman Garlichs

The development of the extensive Southeast Missouri deposits greatly preceded that of the Iowa and Wisconsin deposits. It began about 1720 at Mine La Motte and other localities, and has continued uninterruptedly to the present time. It is estimated that this district produced about 184,000 tons of lead in the year 1916, having a value of approximately $25,000,000. This is a new high record. The Southwest Missouri orebodies were hardly known before 1845 and were not extensively developed until 1870. Lead production in Southwest Missouri for 1916 was 38,788 tons. Oklahoma produced 14,399 tons and Kansas 2,345 tons, making a total for Southwest Missouri, and the adjoining districts in Oklahoma and Kansas, of 55,532 tons of lead. The first metallurgists in Missouri were the Indian and the hunter. The early settler learned to procure his bullets either by melting ore in his camp fire or by throwing pieces of galena on an old stump and depending on the usual roasting and reaction method to obtain the metal. A little more refined method was to arrange two flat stones in the form of a, V. Wood and galena were placed in the furnace and ignited.. As heat developed, some of the metal was extracted and molded into bullets. The next step in utilizing the lead ores was the log hearth. All lead in Missouri was smelted in this crude affair before 1820. It was built on sloping ground, and consisted of a hearth of stone, surrounded on the front and two sides by a stone wall. The wall was 7 it. high in front. The top and rear end were left open and in front an arch or opening was made, forming the eye of the furnace. In front of this a pit was dug in the ground, to receive the molten metal. Large logs were rolled in at the back and made to rest upon ledges formed inside, to raise them from the hearth and to give a draft. These logs filled the entire width of the furnace. Small split logs were then set up around the two sides and the front, and the ore was piled on until the furnace was full. Finally, the mass was covered with logs and fuel until the ore was completely surrounded. A gentle heat was started, which was raised very gradually. After 12 hr. the heat was increased and continued for 12 hr. more, 24 hr. being required for each smelting. The furnace treated about 5,000 lb. per 24 hr. and the ore yielded about 50 per cent. of its metallic lead. A considerable quantity of the ore was not desulphurized, and fell between

Jan 1, 1918

By N. J. Grant, B. C. Giessen, M. Morris

The slvuclures of alloys in the system In-Bi have been investigated after (levy vapid queuching from the mell (splat cooling) to -190°C. Tuo-phase fields could be suppressed over most of the tota1 concentvalion range; five melastable phases a1, a2, ?, ?1. and ß exisl, of which four have simple elementtike structures. For example, ß has the A5 struclure of white till. Crystal-tographic data and coordination numbers for these phases aye given, as well as addilional information on the equilibrium diagram, where a new phase In5Bi3 teas found. THE phase diagrams of binary combinations of B elements from the groups B2, B3, B4, and B5 are usually of a simple type and show considerably fewer intermediate phases than combinations of these elements with A metals, transition metals, or the B1 noble metals. Aside from phases such as the typical B3-B5 compounds with the ZnS-B3 structure type, few stoichiometric phases are known; among them are In2Bi and 1nBi.l Nonstoichiometric phases include ß (In-Sn), a1, (In-Pb), and the tin-rich ? phases in the In-Sn, Cd-Sn, and Hg-Sn systems.' The first two phases have a close relationship to indium, while the latter group are known to be structurally related to a Sn. Thus, there are generally no immediate experimental data for the correlation of the crystal structures of B metals and B metal combinations with certain monotonously varying parameters, such as the valence electron concentration (VEC), or the average atomic size. Such parameters have been recognized as structure determining, e.g., for the transition 4 Sn — y phase,2,3 where nonintegral valence electron concentrations are encountered. It should be possible to extend such correlations to other binary systems, if nonequilibrium single-phase alloys could be produced over broad valence electron concentration ranges, and if their crystal structures could be regarded as being imposed by their electronic states. It has been shown in the case of the tin-rich 7 phases3 that alloy phases with typical crystal structures can be produced in B metal alloys rapidly quenched from the liquid to -190°C by the splat cooling technique due to Duwez,3-8 and reviewed in Refs. 5 and 6. The In-Bi system was selected because it extends over a valence electron concentration range in which several metastable, nonstoichiometrir phases could be expected to occur. Further, the low melting points of the known intermediate phases, In2Bi and InBi, Fig. 1(a), indicated low binding energies and thus possibly low driving forces for the formation Of the equilibrium structures. EXPERIMENTAL TECHNIQUES AND RESULTS The preparation of the quenched foils followed the practice described in Ref. 3. Master alloys were produced by melting of indium (99.97+) and bismuth (99.99+) in evacuated Vycor capsules or in an inert gas arc furnace; quantities of 20 mg were splat-cooled onto copper and silver substrates held at - 190°C; and crystal structures and lattice parameters were determined on a GE XRD-5 diffractometer using Cu Ka, radiation at -190°C and at room temperature. The duplication of each run with both substrates permitted the elimination of overlap of the substrate diffraction pattern and that of the investigated substance. The XRD patterns were usually taken from sin2 " = 0.03 to 0.35; the substrate was used as a means of internal calibration. The fractional accuracy of the lattice parameters of metastable phases is approximately 10-3 In the following, all percentages are in atomic percent. The stable and metastable phases found after rapid quenching to -190°C are listed in Table I, together with estimated concentration ranges and crystallographic data. The investigated alloys and phases present in them are given in Table 11. Except for the region between indium with 33 and 50 pct Bi, where a revision of the equilibrium phase diagram became necessary to include a new stoichiometric phase In5Bi3, the diffraction patterns taken after heating to room temperature agreed with those expected from the phase diagram, Fig. l(a). This suggests that the new nonstoichiometric phases are not stable at room temperature, thus following the observations made for the y phases based on tin.3 Results concerning the equilibrium phases In2Bi and In5Bi3 and the revision of the equilibrium phase diagram which is made necessary by the inclusion of In5Bi3 will be treated first. The Crystal Structure of In2Bi. Makarov7 had identified In2Bi as belonging to the AlB2-C32 type ; however, in a later paper the structure was revised.8 In2Bio was found to be of the Ni21n-B8ß type,21 with a = 5.496A, c = 6.57.9A, and N = 6 atoms per unit cell.9 This crystal structure was confirmed in the present work; definite evidence for the doubling of the c axis, as compared to the A1B2-C32 type, and for the proposed order structure, was found in the powder pattern. The observed lattice parameters agree well with those of Ref. 8; those measured at -190°C are given- in Table I. The Crystal Structure of In5Bi3. A new, equilibrium, intermediate phase was identified in slow-cooled, as-cast alloys; it was identified as In5Bi3. The structure has been worked out by Giessen and Grant;" lattice parameters at -190°C are given in Table I. This phase is probably identical with "In3Bi2", produced by vapor deposition techniques, but not recognized as an equilibrium phase by Palatnik et a1.11 After completion of the present work, the existence of In5Bi3 has also been demonstrated by superconductivity measurements on In-Bi alloys.20

Jan 1, 1968

By C. B. Alcock, M. G. Hocking

The kinetics of the growth of sulfate layers on blocks of Cu20 has been studied between 688" and 800°C. At 788c the over-all weight increases followed the parabolic rate law, with k = 216 30 sq mg per cm4 min, using a 1:1 entering gas mixture of SO2:02. At 757 after an initial linear period the parabolic rate law was followed, with k = 57.8 10 sq mg per cm4 min. The gas mixture required a platinum catalyst to produce the maximum SOs pressure. The sulfation rate at 757' and 788c was high until all the CUZO had changed into a core of CuO, when it became very much lower. At 757°C marker experiments showed that the sulfates (CuO. CuS04 and CuSO4) grow outwards from the oxide surface, but CuO. CuS04 mixed with CuO was found beneath the markers. At 688 markers were found at the CuO. CuS04/CuS04 interface. Visual observations at 757 showed a liquid eutectic moving beneath the sulfate surface. causing very rapid sulfation and extruding to solidify as thick 'needles" up to 8 mm long. Marker experiments have shown conclusi.cely that CuO grows on Cu20 in air by the outward migration of copper. ThermODYNAMIC measurements on the sulfate and basic sulfate of copper, reported recently,' are followed here by measurements of the rates of formation of these sulfates on cuprous oxide, in order to provide a background for the understanding of roasting processes. peretti2 and McCabe and Morgan3 have shown that the phase Cu20 lies next to Cu2S when the sulfide is roasted under conditions where sulfates will be formed. Since this would present too complex a system for kinetic study, the present work4 has been based on the kinetics of sulfation of the oxide phase. The results of a study of the kinetics and mechanism of the sulfation of cobaltous oxide have recently been reported.4,5 Results on the sulfation of cuprous oxide have also appeared,6 but are at variance with those to be reported here. This seems to be because the previous workers6 did not obtain constant SO3 pressures from the SO2/O2 sulfating gas mixtures used, as will be discussed later. Wadsworth, Leiter, Porter, and Lewis found that the sulfates produced when SO2/O2 mixtures are passed over Cu20 at red heats are CuS04 and the basic sulfate CuO.CuSO4. When the decomposition of CuS04 was followed' by differential thermal analysis, a stage corresponding to CuO .CuS04 was reached before decomposition into the oxide occurred. APPARATUS The vertical furnace used for quantitative studies was a conventional resistance-heated and propor- tionately controlled furnace with an even zone (1 deg) which was 4 cm long. The gas mixtures of SO2 and 02 were made by mixing measured flows from capillary flowmeters and drying over phosphorous pentoxide before passing them into the Pyrex top-piece of the furnace assembly. A mullite tube of bore 2% cm was used as reaction vessel and at the bottom of this tube was a trap for SO3, which condensed as a colorless liquid. The residual gas was led into a fume cupboard for ventilation to the atmosphere. The furnace used for visual observations of needle growth and other phenomena consisted of a transparent silica reaction tube of bore 3 cm, having an optically flat silica window sealed into its wall. The reaction tube was clamped vertically, inside a mullite tube of bore 4 cm and length 42 cm which was heated by Kanthal windings. The window in the silica reaction tube was arranged opposite a hole 3 cm wide and 2 cm tall which had been cut in the wall of the mullite winding tube 19 cm from its top end. A binocular microscope was used to look through the window, illumination being provided by a focusing microscope-lamp offset sideways and downwards. The oxide block to be sulfated was suspended from the removable Pyrex top-piece to the reaction tube, by means of which it was inserted at the start of the experiment, and through which entered the S02/02 gas mixture. The variation of the temperature inside the reaction tube with horizontal and vertical displacement from the window was investigated, and, while it can be said that the temperature was constant 1 deg, its value at a given position was known only to 12 deg. In both furnaces, the oxide block to be sulfated was suspended from a small silica hook immediately downstream from a platinum thimble used as catalyst, which was held on a long silica rod from a Pyrex spring attached to the removable Pyrex top-piece of the furnace. The spring was only included in the quantitative sulfating furnace; its sensitivity was 1.4 cm per g and it was found to be free from creep in a 16-hr test when carrying the working load. The extensions were followed by a cathetometer which could be read to 0.001 cm, corresponding to i0.7 mg. Extensions of the spring were measured on adding weights between 5 and 500 mg to the working load (thimble + oxide block), and Hooke's law was closely obeyed. SAMPLES Cuprous oxide blocks were prepared by oxidizing Johnson Matthey "Specpure" copper sheet of 1 mm thickness (impurities: iron at 3 ppm, silver at 2 ppm, lead, nickel, sodium, and silicon each at 1 ppm, calcium, chromium, magnesium, and manganese each at less than 1 ppm) for 6 days in air at 1000°C in a re-crystallized alumina boat. It has been reported7 that when Cu20 is prepared by heating copper sheet for 1 day the central plane of the blocks produced consists

Jan 1, 1967

By A. C. Fraker, H. H. Stadelmaier

In the systems Mo-Fe-C, Mo-Co-C, and Mo-Ni-C the phase boumdaries of the cubic Ij carbides were determined at 1000°C and the variation of the lattice parameter with composition was measured. In the system Mo-Fe-C a single ternary carbide, Mo3Fe3C, was observed. In Mo-Co-C and Mo-Ni-C two new n phases of the type M12C were found: Mo6C06C and Mo6Ni6C. At 1000°C they coexist with the M6C phases MO3-4,CO3-2C and Mo3-4Ni3-2C, respectively. Mo3C03C and Mo4C02C are not separate phases but parts of a single M6C phase region, and the same is true of Mo3Ni3C and Mo4Ni2C. THE n carbide structure type has been known since it was determined by westgren' in 1933, yet reliable information on phase compositions of this type is ap-parently scarce. The following selection of suggested formulas illustrates the extent of the confusion con-cerning compositions: W2-3Fe4-3C,1W6Fe6C,2 W2_4Fe4-2C,3 (W,Fe)6C,4 Tj.-W4Co2C and nz-W6Co6C,5 n1-W3Co3C and n2-w4Co2C,5 W3C03C and w6Co3C2,7 MO3Fe3C,8n2-M04Fe2C,9 (MO,Fe)6C,4 MO3CO3C,8 n2-MO4CO2C.9 Many of these compositions obviously overlap, some disagree in carbon content. The use of the designations 77, and n2 is not consistent. There is uncertainty whether two carbides M6C coexist or whether there is only one M6C with a wide range of substitution of one metal for the other. The existence of M12C seems questionable altogether if one ignores the results of a neutron diffraction study of W6Fe6C supporting that particular metal to carbon ratio.'' More reliable information is available for the iso-typic oxides studied by Nevitt,11 Nevitt, Downey, and orris," and Meussner and carpenter.l3 That n car-bides and oxides are structurally identical, particularly with respect to the location of the metalloid atom, has been demonstrated clearly by Stadelmaier and Meussner14 and also by Parthe, Jeitschko, and sadagopan.15 Therefore the present paper, which attempts to establish the phase boundaries of n defini-tively for three carbon systems, has implications for the whole family of filled Ti2Ni types. EXPERIMENTAL PROCEDURE The samples were prepared in two stages. Pellets weighing 3 g were compacted from powders of the elements and presintered in a closed tube of fused silica that had been evacuated to lO-4 mm Hg. This was fol- lowed by arc melting the sintered samples under argon. Remelting the button once was sufficient to produce homogeneous specimens. The metal powders had the following minimum purities in wt pct: molybdenum 99.5, iron 99.0, cobalt 99.7, nickel 99.9, graphite 99.94. Carbon determinations by the gas volumetric method were obtained from thirty-one of the one hundred samples in the annealed condition. They were limited to compositions that were critical to placing a boundary line in the isothermal sections. The carbon losses were comparatively low and averaged less than 1.5 at. pct. For annealing, the samples were sealed in tubes of fused silica evacuated to 10-4 mm Hg. The annealing schedule was 20 hr at 1250°C followed by 150 hr at 1000" ±10°C. For the water quench the silica glass tubes were broken mechanically to speed the quench. The specimens were examined in the as-cast and annealed condition by standard metallographic techniques. Some of the as-cast and all annealed samples were studied by X-ray diffraction using a 114.6-mm powder camera with filtered Co radiation. The lattice parameters of the n phases were determined by graphic extrapolation to 8 = 90 deg on a plot against sin2 8. It is estimated that the maximum error in the lattice parameters is ±0.002A.. The phase fields were determined by the disappearing phase method, relying on X-rays for phase identification and on microscopy to detect small amounts of a phase. RESULTS Pertinent information on the binary side diagrams has been summarized by Hansen and Anderko16 and Elliott17 and will not be discussed further. Photomicrographs documenting the equilibria at 1000°C have been selected to avoid redundancy due to the similarity of the three systems. Thus there will be only one picture showing an equilibrium with Mo2C, and so forth. Molybdenum-Iron-Carbon. An isothermal section of most of the Mo-Fe-C system at 1000°C is shown in Fig. 1. No attempt has been made to distinguish between a and y iron or to determine their phase boundaries. A conspicuous absence is the phase Fe2 1Mo2C6 with the Cr23C6 structure, supporting the suggestion by Krainer4 that it decomposes by annealing at high temperatures. The only ternary phase is found in a narrow range around Mo3Fe3C (MO43Fe43C14). A microstructure with a two-phase mixture of Mo + n is seen in Fig. 2. The variation of the lattice parameter of n with the estimated composition is shown in in Table IA. The phase diagram does not agree with the results of Kuo9 who reported n2-Mo4Fe2C. Our lattice parameter Varies from 11.095A at the iron-rich end to 11.140Å at the molybdenum-richoend of the phase field whereas Kuo9reported 11.26Å for Mo4Fe2C. We are in agreement with the older work

Jan 1, 1970

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core. For this reason neither analysis of core recovered (amounting to 37.5 pct of approximately 10,000 ft of hole drilled) nor sludge analysis considered separately would give correct assays. The following equation has been adopted: actual weight theoretical actual weight core — theoretical weight sludge Drilling is done un.der contract. Collection of core and sludge samples together with double-checked assaying is done by the company. Microscopic, specto-graphic, and X-ray examination of rock specimen saved is carried on at the University of Utah. A fund, set up by Kennecott Copper Corp. at the University, has encouraged cooperation in solution of mining research problems. The drill used is a Longyear hydraulic powered by a Waukesha 26-hp gasoline motor. A pump, Bean Royal type, is powered by a 7-hp Briggs-Stratton gasoline motor. Weight of unit is 4000 lb and is placed under tripod of 35-ft wood poles within a 25-ft sq area. The following results have been attained: Avg depth of hole drilled, ft 1215 Avg drilling depth per 8-hr shift, ft 10.11 Core Size, In. Avg depth of NC casing, ft 21 2 11/16 Avg depth of NX casing and size core, ft 178 2Ye Avg depth of BX casing and size core. ft 427 1% Avg depth of AX casing and size core, ft 585 1 5/32 Through the use of metal sludge collection boxes, 12 ft long, 18 in. wide, and varying from 12 to 24 in. in depth, sludge recovery has been maintained above 95 pct. Samples, both core and sludge, are assembled for each sample interval and averaged in 50-ft bench increments. Determinations for mineral values are made at both the mine and mill. Assay results varying in excess of 0.05 of copper assay values are rerun to determine correct assay. The organization chart for the Kennecott Utah copper mine is shown in Fig. 1. The Utah mine requires approximately four men per shovel shift to drill solid rock, properly slope banks, and trim loose material from the top of bench sections in making them safe for men and equipment working below. During recent years, while the mine has been worked day and night, powder crews have worked day shift only. Series of toe holes drilled back of advancing electric shovels comprise approximately 70 pct of all mine drilling. These holes, spaced about 21 ft apart, have a depth of 24 ft. Of the footage drilled, 25 pct is for vertical holes either 24 or 28 ft deep, spaced about 28 ft centers along the top rim of banks. This combination of toe holes, using an Ingersol-Rand DA 35 drifter mounted on a tripod and the vertical holes using an Ingersol-Rand Type X71 wagon drill is standard practice above the pit elevation. These benches are 68 ft high, with a few exceptions, while benches below the pit elevation are 50 ft high and require very little drilling other than toe holes. Five benches, varying in height from 80 to 100 ft, require a combination of toe holes plus vertical holes and bank holes or electric churn drill holes. Bank holes using the DA 35 drifter and electric churn drill using Bucyrus-Erie Model 29T, account for 5 pct of feet drilled. Until the past year, drilling of fill material on upper levels was accomplished by churn drill method. Wagon drills are now used for fills because they can be set up and drill a hole an hour. It is necessary to run an air drill 7 hr for each 8-hr shovel shift. Average drilling rates during the past year were as follows: DA 35 drifter using 2 and 3-man crews — 4.36 ft per man hr X 71 wagon drill using 2-man crews — 6.45 ft per man hr DA 35 drifter bank holes, 3-man crews — 2.53 ft per man hr Bucyrus-Erie churn drill, 2-man crews —1.9 ft per man hr Total mine avg — 4.57 ft per man hr Mine practice assigns powdermen to job locations where they perform all drilling and blasting work within the area. Almost as much labor is consumed in springing and loading holes as in drilling them. Average powder consumption is approximately as follows: Per Electric Shovel Shift, Lb To spring and chamber holes — 50 Blasting — 500 Trim banks — 50 One fifth of all powder used is Hercules hercomite bag powder while four fifths is Hercules gelamite 65 pct strength in 1 Vi in. x 8 in. sticks. Frozen material, overhanging from top banks, creates a winter hazard to shovel operators and equipment. Experiments are being conducted with a high-speed Hercules powder in an effort to eliminate this condition. Banks are trimmed by powdermen who lodge 8 or 10 bundles of powder, containing 15 or 20 sticks each, across the upper surfaces of each working face. To each bundle of powder is attached a primer with 6-ft fuse which is lighted as the powdermen, who have lowered themselves down the bank on ropes, ascend to safety. Three angle dozers, D8 caterpillars, make grades and move supplies for the average 31 drilling machines working daily. Compressed air is furnished by three Ingersol-Rand compressors, each with rated capacity of 3750 cu ft of air per min at 100 lb reservoir pressure.

Jan 1, 1952

By C. H. Curtis

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

Jan 1, 1951

By C. H. Curtis

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

Jan 1, 1951

By E. H. Boswell, E. M. Cushing

The Mississippi embayment east of the Mississippi River is part of a vast geologic and hydrologic province. The abundance of water available is largely responsible for the agricultural history of the area and is perhaps the key factor in the rapidly accelerating trend toward industrialization. The multiple artesian aquifer systems that underlie the region and the alluvial aquifer along the Mississippi River are virtually undeveloped sources of fresh water in that they are yielding only a fraction of the water available. The potential yield of the 12 aquifer systems is estimated to be about 25,000 mgd (million gpd) of which only about 600 mgd is presently being withdrawn. The amount of water in storage in these systems is about 1600 trillion gal. Ground water having a dissolved-solids content of less than 500 ppm is generally available at depths of less than 1000 ft, and water having a dissolved-solids content of less than 1000 ppm is available in some places to depths of more than 3000 ft. Iron is the most common troublesome chemical constituent in the ground water. Water in varying amounts is also available from streams within the region. The amount of water which originates within the region and which leaves it as streamflow during a year averages about 50 million acre-ft (about 45,000 mgd). An additional 375 million acre-ft (about 340,000 mgd), which originated outside the region, leaves the region as streamflow during an average year. Some streams have considerable flow during periods of no precipitation and the water during these periods generally has a low dissolved-solids content. Large quantities of water from streams and waterbearing sands (aquifers) are available in the Mississippi embayment east of the Mississippi River. Most of the region is underlain by aquifer systems that will yield large quantities of water to wells, making ground water the most readily available source of fresh water. About 90% of the water used in the region is obtained from wells. Streams have been used primarily to dispose of, and dilute, wastes that have been dumped into them. The Mississippi, Ohio and Tennessee rivers, and the lower parts of the Tombigee and Yazoo rivers also serve as a means of transportation. This nominally water-rich region receives from 48 to 56 in. of precipitation in average years and is crossed by several large rivers. The largest, the Mississippi River, is the western boundary of the region discussed in this paper. The economy of the region is basically agricultural, but industry has become increasingly important since World War 11. The diversification and mechanization of agriculture have resulted in a surplus of manpower, most of which is being absorbed by the expansion of industry. Further industrial expansion probably will occur because of the available manpower, natural resources and economical transportation. Increasing demands on available water supplies will be made and the future economic growth of the region will be largely dependent upon wise utilization and management of the water resources.

Jan 1, 1969

Jan 1, 1965

By P. D. White, J. T. Moss, J. S. McNiel

Results of experimental work on the in situ combustion process were first published in this country in 1953' when Kuhn and Koch described results of a three-well test in Jefferson County, Okla. Shortly there-after Grant and Szasz2 described field studies in Nowata County, Okla., in which several modifications of the basic combustion process were investigated by injection of various oxidizing gas mixtures other than air. In the past five years speculation concerning this method of oil recovery has continued, with a number of interesting papers on various aspects of in situ combustion appearing in the literature3,4,5,6,7. However, only one of these' provided actual engineering information on pattern-type field experiments. The primary objectives of the Oklahoma three-well test were to check operation of specially designed electrical ignition equipment and evaluate problems associated with initiation of combustion in a natural oil sand reservoir. Immediately following successful completion of this test in the Pontotoc sand, plans were activated for a second field experiment to be conducted in the same sand a short distance away. The objective of this second field test1 as to establish a relationship between field observations and information derived from laboratory experimentation and computations. Secondary objectives of the test were to develop operating tech- niques and define operating problems, and to obtain information that would assist in the economic and engineering appraisals of the process. Careful planning was required since this test was designed with the hope of providing basic information on the characteristics and operation of a combustion drive. It was necessary to include test facilities which would not only permit precise control and measurement of the process variables but would also provide sufficient flexibility of control to insure that variables in the combustion process could be studied during the course of the test under a wide range of conditions. In order to accomplish these objectives elaborate instrumentation and several control wells were included in the design of the test installation. This type installation was felt to be well suited for an experimental study, but was in no respect intended as a counterpart of a commercial-scale operation. PRELIMINARY LABORATORY WORK An important objective of the field experiment, as stated previously, has been to obtain reliable data on operating variables which might be compared with similar data obtained in laboratory combustion tests. The value of successful correlations between quantities that can be measured in the laboratory and those obtained in a full-scale commercial venture has been well established for other industrial processes. The difficulty of following a process underground should make such correlations even more essential here. Also. approaching the field test phase of development of in situ combustion, it was considered desirable to precede design of test equipment with laboratory combustion studies under a variety of conditions to establish likely ranges of variables. These included temperature, air injection rate, frontal advance rate and produced gas composition.

By H. L. Miaw, A. M. Gaudin, D. W. Fuerstenau

A quantitative method for evaluating density of slime coatings has been developed and applied to formation of iron oxide slime coatings on quartz and on corundum. Slime coating density is related to flotation recovery and to properties of the electrical double layer at mineral surfaces. IN spite of considerable study, the nature of slime coatings in flotation is still not completely understood. However, phenomena that control floc-culation and dispersion of colloidal systems are now interpreted in terms of electrical double layers.",' Colloids will flocculate even though the particles carry the same charge, but when two dispersed colloids of opposite charge are mixed. flocculation ensues even more rapidly. Analogous phenomena seem responsible for slime coating on minerals in flotation systems. To shed new light on the issue it was considered desirable to devise a tool that would measure slime coatings. The usual method has been to examine photographs of presumably representative surfaces, after. an intervening washing in running water. But such an examination is non-quantitative, and the changes in chemical environment represented by the washing may have removed the slime coatings that existed during flotation. To provide the possibility for a correlation with electrical double layer phenomena, the experiments were made with quartz, corundum, and iron oxide, for which elec-trokinetic data are available.'-" Since the properties of the electrical double layers of these oxides depend on the pH of the solution, pH will be the most important variable. Furthermore, because the zero point of charge for silica, iron oxide, and alumina occurs at pH 3.7, 8, and 9.5, respectively, the relative charge on the slime and mineral can be changed over wide ranges. Because of its practical importance in quartz flotation, iron oxide powder was chosen as the slime. Also, its uniform composition, contrasting color, and ease of analysis make experimentation simple. In this investigation, correlation was sought between the effect on flotation recovery of slime addition and slime coating density. Materials and Methods Details are available in a thesis by Miaw." The quartz was selected from pure Brazilian crystalline quartz, crushed, sized (65/150 mesh) acid-cleaned, washed, and stored in conductivity water. The corundum (synthetic alumina of high purity prepared by Linde Co., Chicago) was crushed, sized (65/100 mesh) deslimed, boiled in nitric acid, washed, and stored in conductivity water. The iron oxide slime, reagent-grade ferric oxide (—400 mesh) from Allied Chemical & Dye Corp., was heated to 160°C, cooled, and stored in a vacuum desiccator. In the electron microscope the slime particles appeared to be about 0.3µ in size, but their large specific surface (87,900 cm per g, by gas-adsorption method, using krypton) suggests that the actual particles are smaller than 0.3r. Had they been perfect nonporous spheres, they would have had a diameter of 0.13µ.

Jan 1, 1959

By G. G. Brown

Enthalpy-entropy diagrams are presented for natural gases of 0.6, 0.7, 0.8, 0.9, and 1.0 gravity over the pressure range of 5 to 10,000 Ib. per sq. in. and temperature range of 32º to 700°F. The charts indicate directly the work requirement and temperature rise for adiabatic compression or temperature change for free emansion of natural eases. Computation and Uses of Charts The Mollier diagram (Fig. 4), in which the enthalpy (heat content) is plotted against the entropy with lines of constant temperature, pressure,, and in some cases volume, has been found most convenient when dealing with the compression, expansion, and flow of fluids. In dealing with the flow of fluids, the sum of the increase in the enthalpy, AH, plus the increase in kinetic energy, AM —, plus the increase in potential energy, ?MZ, representing the total increase in energy of the fluid in flow, is equal to the sum of the heat q, and work added, —w, to the fluid while flowing between the entrance and exit of the flow system. AH + AM— + ?MZ = q-w [1] In cases where there is no significant change in potential energy or in kinetic energy (velocity), it follows that the increase in enthalpy is equal to the total energy supplied to the fluid. Under such conditions' the changes in the property of the fluid as it flows through a throttling valve, choke, or any other similar arrangement, may be read directly from the enthalpy-entropy diagram by following a horizolltal line between the known pressures' When compressing Or expanding a gas by means of a Or or engine in which no heat is added to or subtracted from the gas, but only work done, the changes in the properties of the gas may be determined along a vertical line of constant entropy between the entering and exit pressures. The power required for the compression of the gas may be readily determined by converting the increase in enthalpy into the desired units. The enthalpy-entropy diagram for natural gases is to the gas engineer what the steam diagram is to the steam-power engineer. For this reason it would be extremely convenient if a reasonably satisfactory enthalpy-entropy diagram could be prepared as a function of the gravity of the gas. A careful study of the known properties of natural gas indicate that this is possible. The effect of temperature upon the enthalpy at constant pressure is expressed as the "heat capacity" or "specific heat" of the gas. The best available data for natural and petroleum refinery gases indicates the relationship shown in Fig. I. From this it is clear that the specific heat of natural gases is a function only of the gas gravity and the temperature at atmospheric pressure. The effect of pressure on the enthalpy of natural gases is dependent upon the pressure-volume-temperature relationships,

Jan 1, 1945

By Michael Perch, Charles C. Russell

Nearly 7 5 per cent of the total coke production in the United States in 1940 was consumed in blast furnaces. In 1939 the percentage was 69.9, and in 1938 it was 61.3. To produce a net ton of pig iron 1757 lb. of coke was required in 1940 and 1760 lb. in 1939. These figures indicate how dependent the production of pig iron is upon the production of coke. The rate of iron production is in some measure influenced by the physical and chemical characteristics of the coke. Blast-furnace operators believe that the physical properties of coke profoundly influence the operation of the furnace. Although it is true that coke of a wide variety of physical properties is used in American blast furnaces, the periodic fluctuation of these properties is one of the causes of irregularity of blast-furnace operation. The purpose of this paper is to present a broad picture of the physical properties of by-product coke that are generally considered to be important by coke-oven and blast-furnace operators. Effects of the kind of coal used, the preparation of the coal, the rate of coking, and other factors will be considered. Of all these variables, the kind of coal used is by far the most important. In this discussion use will be made of the method of coal classification by rank1 standardized by the American Society for Testing Materials. Although this method of classification has been an American standard for several years, it has not been given the attention it deserves, especially in connection with the selection of coal for production of coke. Despite the fact that many producers of blast-furnace coke are limited to some particular source or sources of coal, a good understanding of the principles involved in setting up the standard method of classification may lead to a better appreciation of the causes of changes in coke characteristics. Furthermore, where it is necessary to select new coals as a substitute for the original supply, this method of classification can be of great assistance. The data presented have been gathered from many sources, but no attempt has been made to cover the vast literature. Instead, some of the more important data have been selected to show how coals and their carbonization affect some of the physical properties of coke. It must be emphasized that generalizations concerning the behavior of coking coals are presented and that individual coals may vary widely from the specific behavior depicted. Nevertheless, it is believed that the data illustrate certain typical behaviors. Neither was it possible to cover completely all the factors involved; for example, the macroscopic constituents of coal are believed to have important effects on the physical properties of coke, but since there is little information available for such study, and since vitrain, clarain, and durain are neither

Jan 1, 1943