Search Documents

By Alfred R. Powell

LATE in 1944 a group of petroleum and coal technologists was organized in Wash¬ington under the sponsorship of the Petroleum Administration for War and the U. S. Bureau of Mines. This group, known as the Technical Oil Mission, was placed under the authority of the Technical Industrial Intelligence Committee of the Joint Chiefs of Staff, which was then promoting the organization of other missions to be sent to Germany to obtain technical information regarding various industries in that country. The objective of the Technical Oil Mission was to obtain technical data on German production of synthetic oil from coal and all other commercial practices and research projects connected with the oil industry. The primary purpose was to obtain such data as were immediately essential to assist the war against Germany and Japan, with secondary emphasis on technical information that would assist American industry after the end of the war. Since oil-production facilities in Ger¬many and the occupied countries con¬trolled by Germany were the top targets for strategic bombing, the Technical Oil Mission was the first group under the Technical Industrial Intelligence Com¬mittee to enter Germany. A few members of the group flew to London in February 1945 and the majority of the remainder followed in March. In London they were joined by an equal number of British investigators and the combined, mission was then sent into Germany under the authority of the Combined Intelligence Objectives Subcommittee (CIOS). A sum¬mary of the more important information gained by this combined British and American mission is presented below. GERMAN OIL PRODUCTION The oil available to Germany, both natural and synthetic, and including pro¬duction from occupied countries, reached a peak in late 1943 and early 1944. Pro¬duction at this time was at the rate of 1,200,000 metric tons per month, equiva¬lent to i00,000,000 bbl. per year, or about 7 per cent of the wartime production in the United States. Of this production, 50 per cent was crude oil from Romania and Hungary, 12 per cent was crude oil from Germany itself, including Austria, 9 per cent com¬prised coal-carbonization products, that is, tar and light oil, and only the remaining 29 per cent was synthetic liquid fuel made from coal or tar. Synthetic liquid fuel made in Germany was equivalent therefore to less than 2 per cent of war¬time production of crude oil in the United States. The success of the strategic bombing campaign against German oil-production facilities is demonstrated by the fact that by September 1944 the Ploesti fields of Romania and much of the syn¬thetic oil production had been eliminated, so that the amount of oil available to the German war machine was only 25 per cent of the peak quantity mentioned above. By early 1945, when the combined British and American oil mission was entering

Jan 1, 1946

By K Byrne, W. Hawker

The average grade of copper concentrate is steadily declining, forcing the industry to turn to lower grade and metallurgically challenging orebodies. When faced with a low grade, challenging ore, mining companies are often forced to choose between high recovery at low grade or low recovery at high grade. Hydrometallurgical processing of copper concentrates using metathesis reactions, which replace iron with copper in copper-iron sulphide minerals and increase the copper to sulphur ratio in copper sulphide minerals, provide an opportunity to upgrade copper concentrates and potentially allow high copper recovery for a high grade concentrate. Factors such as concentrate mineralogy, reaction time and temperature, and the concentrations of copper, chloride and acid in solution are reported to affect the reaction extent and mineral phases formed by metathesis. In this paper, a copper concentrate has been studied over a wide range of conditions to investigate the influence of these key parameters on the extent of copper metathesis reactions and product phases formed. Solution and solid phases are assayed with the main mineral phases identified by Quantitative X-Ray Diffraction. INTRODUCTION The minerals present in copper deposits determine the prospective processing options either: pyrometallurgical or hydrometallurgical. For copper oxide ores, copper is typically present as oxides, carbonate, chlorides, silicates and sulphates. These minerals are readily solubilised in acid and are usually treated through hydrometallurgical processes. The typical process route for copper oxides is to leach copper into aqueous solution, purify and upgrade the solution phase through solvent extraction and then recover copper as metal by electrowinning (Davenport, King, Schlesinger, & Biswas, 2002). Copper sulphide deposits may contain a multitude of copper bearing minerals, including chalcopyrite, bornite, enargite, covellite, digenite and chalcocite. The copper ore grade and treatment options are affected by the mineral type and distribution. Sulphide ores are typically treated through comminution and mineral flotation to form concentrates, followed by pyrometallurgical smelting. Hydrometallurgy can also be used to process sulphides depending on the mineralogy and impurity levels (Davenport et al., 2002). The trend of decreasing copper grade with increasing impurity levels are common across the mining industry and pose future challenges to copper processing. This trend has prompted research into hydrometallurgical options for upgrading copper concentrates. One promising hydrometallurgical option is using metathesis reactions. Metathesis reactions, sometimes also called double replacement or displacement reactions, are chemical processes involving the exchange of bonds between two chemical compounds or species. In this case, the metathesis reactions involve copper in solution exchanging with cations in the solid sulphide concentrate. The driving force for copper sulphide metathesis is the relative affinity of the exchanging cation elements for sulphide. Based on sulphide precipitation diagrams, the expected order of affinity for sulphide should be Hg2+ > Ag+ > Cu+ > Cu2+ > Bi3+ > Cd2+ > Pb2+ > Sn2+ > Zn2+ > Co2+ > Ni2+ > Fe2+ > Mn2+ (Monhemius, 1977). Hence, copper metathesis reactions are capable of increasing copper content and removing impurities such as iron, nickel, zinc and lead from copper concentrates.

Jan 1, 2019

By M. E. Wadsworth, I. B. Cutler

Coprecipitation of anionic and cationic polyelectrolytes has been applied to floccula-tion of several mineral systems. Results obtained in a study of the flocculation of kaolinite and hematite suspensions of polycationic and polyanionic electrolytes are presented. Greatly increased settling rates were observed following precipitation of positive and negative polyelectrolytes on the surface of finely divided minerals in aqueous suspension. The ratios of polycationic to polyanionic electrolytes required to produce maximum sedimentation have been shown to correspond closely with the equivalence points obtained by light scattering studies of systems containing the positive and negative polyelectrolytes by themselves. DISPERSION of colloids has been of interest to surface chemists for many years. Flocculation studies have been employed to evaluate the limits of colloid stability. In water suspensions, colloids become unstable and flocculate when the ( potential, the potential of the kinetic solvation sheath, falls to a value at which van der Waals forces may act to draw the particles together. This may happen through an adjustment of the pH of the suspension or by addition of a salt. Optimum conditions for rapid formation of large flocs are found in the region of minimum ( potential. In addition to these well known considerations, colloids in water suspensions are known to be flocculated by certain large organic macromolecules. These polyelectrolytes in some cases minimize the i potential and in other instances increase the attractive forces by supplying sites for hydrogen bonding.' Use of organic polyelectrolytes in flocculation of mineral colloids has increased rapidly in recent years. These materials are now recognized as indispensable aids to fluid-solid separations in many extractive metallurgical operations. They have been found to increase both the settling rate in thickeners and the filtration rate of filters. In the elucidation of the characteristics of these organic reagents as floc-culants in this laboratory, it was discovered that interaction of certain polyelectrolytes had a surprisingly good effect on flocculation of mineral colloids. Experimental Procedure: The anionic flocculants used in this study were Lytron 886 and 887, produced by Monsanto Chemical Co. Both reagents are copolymers of vinyl acetate and maleic anhydride. Lime is added to Lytron 886, whereas Lytron 887 is all organic. These reagents possess two carboxyl groups for each acetate group and are therefore anionic in character. The cationic reagents used were the Peter Cooper Co. 1-X and 2-X glues. The structures are not definitely known, but the re- agents are amines and are cationic in character. The glues were prepared in the chrome-alum form. Both types of reagents may be classified as polyelectrolytes in that they are high molecular weight polymers and ionize in water. All data presented are in terms of sedimentation rate, measured in 250-ml glass-stoppered graduated cylinders. With suitable precautions such measurements are fairly reproducible and provide a sensitive means for selecting important parameters such as pH, reagent concentration, and floc size and density.2 Agitation was standardized with five complete end over end cycles of the suspensions in the stoppered graduated cylinders.

Jan 1, 1957

By L. E. Young

LOADING machines may be classified in several ways: (1) Machines which cut or break down and load .the coal; (2) machines which simply load the coal; (3) devices which load and transport the coal; (4) de-vices which elevate; (5) devices which transport the coal after it has been shoveled onto the device. There are many types of mechanical loading devices, Bull. 17, Coal Mining Investigations, under the auspices of Carnegie Institute' of Technology, U. S. Bureau of Mines, and the Advisory Board, listing 31 exclusive. of -the large shovels used in thick coal seams. Mechanical loading of coal is not in its infancy; I think some of the early patents go back to 1890. One large coal company, properly called the early leader in this field of mechani-zation, has about 26 large machines at work. This company spent about thirty years in developing a ma-chine, and at the present time is loading approximately two million tons of coal a year in its mines. The chair-man directed a question at the previous speaker; inquir-ing if real progress was being made in mechanical load-ing. The experience of the company just referred to is in itself an adequate answer. A review of a list of mines using mechanical loading devices shows that there are 28 mines which we may say are 100 per cent mechanical. There are four more mines which have been 100 per cent mechanical, but which are idle at this time on account of labor condi-tions. There are at least 25 more at which the tonnage is more than 50 per cent mechanical. In speaking of a 100 per cent mine I do not mean that every pit car of coal is loaded by machine, because when complete, re-covery of the coal is attempted it is not practical to load all the coal at, the break line. In order to load the last 5 to 10 per cent of coal with a minimum of gob material and in order to keep the. working places along the break line timbered in a safe manner it is generally desirable to use a few hand loaders. At a certain mine operat-ing on the block system 90 percent of the coal is being loaded by what is properly a mechanical loading: device, and about 10 per cent -is being loaded by hand. This 10 percent is right off the break line, out of what might be called, the screen pillar. The mines referred to are scattered over the country, . in seven different' States, and the equipment used is produced by ten different machinery companies. As to size, these mines vary from small mines of 500 tons mining capacity-up. In one instance a company with one loading machine is producing an, average of 450 tons per day. One of the large mines that is idle pro-duced at the. rate of 4300 tons per day, of which 97 per cent was actually loaded by machinery.

Jan 8, 1928

By F. M. Van Tuyl

The oil industry had its inception in the United States in 1859 when the first oil well was drilled by Col. Drake at Titusville, Pennsylvania. From, a production of 2000 barrels for that year there was a slow but steady increase, with some fluctuations, up to 1901 for which year the output was approximately 69,389,194 barrels. Since that time the production has increased with leaps and bounds due to the rapid rise of the automotive industry, the widening use of oil as a fuel under boilers and the development of the. Diesel engine. The production for 1921 is estimated to have been 469,639,000 barrels. During the same year more than 125 millions of barrels were imported from Mexico. The oil industry, including the production, transportation, refining and distribution of petroleum and its products, is now the third largest in the country. Present Condition of the Industry During and immediately after the war the demand for petroleum exceeded the domestic production plus the imports from Mexico with the result that our storage oil was heavily drawn upon. Prices for crude exceeded all previous levels in the history of the industry. Pennsylvania crude sold for $6.10 per barrel and Mid-Continent crude for $3.50. Such prices stimulated wildcatting greatly with the result that several important new oil pools were discovered. Then as production increased the depression came which caused a large falling off in consumption. Pennsylvania crude fell to $2.25 per barrel and Mid-Continent to $1.00. Large oil companies such as the Standard, Sinclair and Empire took advantage of the situation, erected additional storage tanks and bought up millions of barrels to hold until better prices returned. Prices have already improved owing to a better market, but the domestic production coupled with the Mexican importations is still in excess of the demand with the result that the surplus is being stored. At the close of the year 1921, the U. S. Geological Survey estimates that there were 183,890,000 barrels of crude in storage in this country, the largest stocks in the history of the industry, yet an amount sufficient to supply the country for only 119 days at the present rate of consumption. Price reductions are again anticipated in some quarters but in view of the probable falling off in the Mexican output in the near, future due to the rapid rate of exhaustion of several of the most important pools and in view of the expected increase in consumption of gasoline and kerosene during the spring and summer months it is more likely that the present price level will be maintained. Present Sources of Domestic Supply Seven major oil fields are recognized by the U. S. Geological Survey in the United States. These are the Appalachian, Lima-Indiana, Illinois, Mid-Continent, Gulf Coast, Rocky Mountain and California. Two of these, Mid-Continent and California, supply about seventy-five per cent of the present total domestic output

Jan 1, 1922

By C. G. Dunn, F. W. Daniels, M. J. Bolton

J. P. Nielsen—The data that Dr. Dunn and his associates have been obtaining are welcome checks on the theoretical aspects of grain boundary energies. With reference to the comments on the validity of the inter-facial tension concept, I should like to say that I have been having considerable difficulty myself in untangling the confusion between "interfacial tension" and "interfacial energy." The confusion however resolved itself for me when I learned that "interfacial tension" for solids, or at least crystals, is used with two distinct meanings. One meaning12. ' "hat is implied for "interfacial tension" is the work necessary to produce or increase the surface of the solid in question by a unit area, temperature and pressure held constant. This comes from the somewhat surprising but simple fact that surface tension and surface energy in the case of liquids are identical quantities. This synonymity has carried over to crystals, where it should be kept in mind that the tension connotation has meaning only by analogy with liquids and that no real stresses are being referred to. The only reason that there is apparently a set of stresses acting at a grain boundary junction is the optical illusion that liquid interfaces are being observed for the grain boundaries. The other meaning3,14,15 for the term has to do with actual states of stress in the surface of a material as would be measured by a system of forces applied to the surface to equate the surface stress state with that in the body of the crystal. J. B. Hess—Since the authors have proposed that the surface tension concept should be discarded in order to deal with the general case of interface energies in solids, their attention should be called to a recent paper by Gurney." The latter has shown, from a thermodynamic argument, that surface tension is a valid concept in solids. In fact, Gurney has rejected the idea that surface tensions are merely mathematical fiction and, instead, has given an interpretation of their physical reality. Finally, he has concluded that, as long as appreciable atomic migration takes place, specific surface energies and surface tensions must be equivalent in solids as well as in liquids. Accordingly, the authors' eqs 2 and 3 can be written with equal accuracy in terms of interface tensions, simply by replacing each E,! with its numerically and dimension-ally equivalent Interfacial tension ?,,. If the interfacial tensions (or specific energies) are independent of interfacial orientation, as is true in liquids, then the general equations that define the geometry at equilibrium of the common junction of three interfaces reduce to eq 1. However, neglect of the effect of boundary orientation is never completely justified theoretically' in dealing with crystalline solids; hence, use of eq 1 in such cases must always be considered as an approximation, though clearly a reasonable one for many purposes.1,2 At the same time, one should not be surprised that the assumption that interface tensions in solids are free from orientation

Jan 1, 1951

By M. Measures, W. A. Napier, C. Pomroy, J. M. Jardine

INTRODUCTION In most uranium mines the concentration of uranium in the ore is low enough that inhalation of dust does not present a radiological problem. Conversely in the mills there is a potential inhalation hazard from inhaled uranium in concentrated form, namely "yellowcake" and other compounds intermediate between the ore and the final product of the mill. It was generally assumed that the solubility of yellowcake placed it in the category of Class D materials. This plus the fact that the uranium is non-enriched led health physicists to assume that the potential health hazards were a result of the uranium's chemical properties rather than its radiological properties. Bioassay programs have thus been limited to routine urine sampling at a weekly or monthly frequency. In most cases when a high reading was obtained, subsequent sampling within one or two days showed a rapid return to low levels, therefore tending to confirm the class D nature of the material. Two developments over the last few years have however suggested that modifications to this simple bioassay program may be necessary and practical. Firstly, a number of studies on the solubility of uranium mill products, from ore to yellow cake, in simulated lung fluid (Eidson A.F., Mewhinney J.A., 1980) have shown that some of the material is not class D, but class W or even Y. This raised the possibility that, while the urine levels rapidly decreased, indicating rapid excretion of the class D component, the W and Y components remained in the lung. Thus an unsuspected lung burden could have been accumulated. The second development, the ready availability of [in vivo] counting facilities with which to make direct measurements of uranium lung burdens, offered a means of confirming the hypothesis of unsuspected lung burdens. A study was therefore initiated to correlate lung burdens with urinary excretion. METHOD A total of 79 subjects from 3 uranium mills participated in the study. They were brought to the Ottawa laboratories of the Radiation Protection Bureau in groups of 4 or 6, over a period of several months. In Ottawa, [in vivo] thorax measurements were performed, and a urine sample collected. Each subject also brought with him a urine sample that had been collected the previous day. All samples were split, for analysis by both R.P.B. and the mill laboratory. [Concentrations of Uranium Dust] Results obtained by mills A, B and C for concentrations of uranium in air at various locations are given in table 1. Samples taken at mill B were obtained using continuous sampling with a midget impinger over a six hour period. Samples taken at mills A and C were 5-minute gravimetric samples. [Subject Selection] The subjects were selected to represent a wide range of work history in terms of both length of employment and type and location of job, as shown in table 2.

Jan 1, 1981

By Richard M. Foose

MOST of the roofing granules produced in Pennsylvania are made by two companies at three plants. The Advance Industrial Supply Co. has three quarries and a mill at Gladhill Station, in southern Adams County. The Funkhouser Company is actively engaged in mining, quarrying and milling at Charmian, about one mile west of Gladhill Station, and at Delta, in southern York County. These three plants produced approximately 100,000 short tons of granules in 1943. The production of roofing granules in 1937 for the whole country was only about 500,000 tons, generally indicating the large growth of the industry since that time. The fundamentals of producing roofing granules were discussed comprehensively by G. W. Josephson in a paper before the Industrial Minerals Division in October 1943 (Mining Technology, March 1944). The purpose of this paper is to describe the actual methods of mining the rock and milling it to make roofing granules, as practiced by the three Pennsylvania plants. MINING AND MILLING Funkhouser Company (Delta) The Funkhouser Company, near Delta, in southern York County, Pennsylvania, operates three joined quarries and a mine in the Peach Bottom slate. The joined quarries extend 1200 ft. in a northeast southwest direction along the strike of the slate cleavage and are about 200 ft. wide. The working face is about 45 ft. high. The quarrying operation is rather simple: Ingersoll-Rand Jackhamers, 1-in. and 1 1/4-in., with hollow drill steel, are used to undercut the quarry face with10-ft. holes. The face is "blown" and allowed to slump on the quarry floor. Loaded by a Marion electric shovel into trucks, the shattered rock is taken to the primary crusher 2500 ft. away. In addition to the normal quarrying operations, a drift into the northeast wall of the middle quarry was begun about 1935 and has advanced about 2000 ft. since then. Rooms have been driven at angles to the drift along choice beds of slate. The drifts and rooms have a roof more than 35 ft. high and the main drift and rooms are all more than 40 ft. wide. There is no cribbing at any place in the mine. Several of the rooms were driven upward on a 5º slope and in some of them there was stoping. Ingersoll-Rand Jackhamers are used, and only wet drilling is practiced. When visited in January 1940, an electric shovel was operating at the head of the drift, scooping up the slate blasted out of the working face. This, too, is trucked to the primary crusher. Some of the better quality slate is saved for making slate shingles. The primary crusher, which is a No. 30 McCulley gyratory, receives the slate from the trucks and crushes it to about 2-in. size (Fig. I). This product travels by a 30-in. belt conveyor to Niagara vibrating screens with 2-in. holes. The minus 2-in. product moves to a storage silo by bucket

Jan 1, 1945

The Candelaria deposit was first discovered by Spanish prospectors in 1863; but its silver ore was not exploited on a large scale until 1873, when the Northern Belle mine was opened. The first Northern Belle mill was built at Belleville, 13 km (8 miles) to the northwest. From 1876 to 1886 the Northern Belle paid over $1 million in dividends each year. In 1882 a 43 km (27 miles) pipeline was built to Candelaria from the White Mountains to the west to furnish water for the mills and town at Candelaria; it was 100 mm (4 in.) in diameter, hand- riveted, and made in England. Sections of this pipe are used. as claim posts today. At its terminus at Metallic City, above Candelaria, a large concrete reservoir was built and called Metallic Lake. It is being used today by Candelaria Partners. Production at Candelaria declined after the mid-1980s, revived for a short period in the early 1890s, and during the 1920s a 300 ton mill operated for three years. Candelaria Partners is owned jointly by Occidental Minerals and Congdon and Carey, with Occidental being the operator. An engineering study done in mid-1979 resulted in engineering for the plant commencing in late that year and heap leaching being started in September 1980. Two million tons of ore per year containing 108 g Ag and 0.28 g Au per mt (3.15 oz Ag and 0.008 oz Au per st) will be mined from the Lucky Hill and Mount Diablo .areas about 1.6 km (1 mile) up Pickhandle gulch from the heap leaching operation. Waste -to -ore ratio is about 2:l. Ore is mined by front-end loaders and hauled by 85 ton trucks to the primary crusher, a 1,070 by 1,780 mm (42 by 70 in. ) gyratory. Primary crushed ore is conveyed to a stockpile of 10,000 tons live capacity from which it is reclaimed by hydraulic reciprocating feeders and conveyed to the crushing and agglomeration facility. This unit is composed of a 1.8 m by 2.6. m (6 ft by 12 ft) double deck screen, a 2.1 m (7 ft) standard cone crusher, and three conveyors. At the conveyor transfer points NaOH and NaCN solution is sprayed on the ore as it falls through a row of 75 mm (3 in.) hanging steel bars. During this process of turning and dividing the ore, the fines agglomerate with the coarse ore to give better percolation for the leach solution on the heaps. Provision was made for the later addition of 2 kg per mt (4 lb per st) of cement to aid in the agglomeration. Agglomerated ore at minus 50 mm (2 in.) size is then conveyed to a 390 mt (430 st) loadout bin where it is transferred to 85 ton rear dump trucks for haulage to the leach pads.

Jan 1, 1981

By Reed W. Hyde

Some information has been published on the sulphur dioxide concentration of gases from D wight-Lloyd machines sintering lead ores but most of this relates to the customary practice in which the charge is relatively high in lead and is sintered in two passes Over the machine. It is thought, therefore, that the following observations on sintering in one pass a low-lead, low-sulphur charge may be of interest. The data were obtained in sintering lead sulphide notation concentrates mixed with minus 20.mesh oxide ore, the mixture carrying 16,8 per cent Pb and 5.3 per cent S. No returns were used because of the limited amount of sulphur available as fuel, but the mixture pelletized readily. to make a satisfactorily permeable bed on the sintering machine. As there was a possibility that the SO z might be utilized, a series of runs was made to determine the concentration ofSO2 in the gases at different points along the wind box, and the effect on this concentration of changes in the wind-box vacuum. Several runs each were made with vacuums (as measured in inches of water at the wind box) of (1 2 to 1 in., ('1 4 to 3% in. and (3) 6 to 5% in. The averages of the determinations from each series are platted On the graph (Fig, 1). The vertical scale shows the percentage by volume of SOz (as determined by the Reich method), and the horizontal scale shows the time e~pressed as percentage of the total sintering time. Sintering was considered to be complete when no SOz could be detected in the gases. In all cases the depth of bed was 4 1/2 in. and the weight of dry charge per square foot of hearth area was 45.2 lb. Other data are given in the table. The volume of gases corresponding to the several points along the wind box at which SO2 determinations were made could not be measured, but has been calculated from the SO2 Content and sulphur elimination on the assumption that the volume is uniform throughout the sintering period. This is not strictly true, as with a given vacuum the volume usually will drop slightly toward the middle of the period and increase toward the end. However, this made is a reasonable approximation. several interesting features are brought out by these data. In general, the published data on SO 2 concentration in sintering under ordinary conditions show a high SO2 concentration in the early stages followed by a rapid decrease in concentration and with relatively little SOz in the gases from the last half or third of the operation. In other words, the greater part of the sulphur elimination is obtained in the first third or half of the period. In these runs, the 6-in. vacuum, which approximates the usual operating condition, gives a curve that follows this customary pattern. Maximum SO2 concentration is obtained 234 min. after

Jan 1, 1944

By J. W. Downey, M. V. Nevitt

MUELLER and Knott' have reported that Ti2Cu has the Si2Mo-type (Cllb) structure. The lattice parameters of Ti2Cu and the positions of the titanium and copper atoms are given by them as follows: ao = 2.944A; co = 10.786A 6 atoms per unit cell: 2 Cu (a) 000; 1/2: 1/2, 1/2 4 Ti (e) 002; 1/2, 1/2, 1/2 + z 00 -2 ; 1/2. 1/2. 1/2- z where z = 0.341 We have found eight other binary compounds which appear to be isostructural with Ti2Cu. All of them involve the combination of a titanium group element with a copper group element or a titanium group element with palladium. The compounds and their lattice parameters are given in Table I. The alloy samples were prepared by arc melting metals having a purity of 99.9 pct or better. They were homogenized for 72 hr at 900°C in evacuated capsules and quenched in tap water. Filings were prepared, and these were reannealed for 4 hr at 700°C, also in evacuated tubes. The lattice parameters were determined by a least squares Bnalysis of the back reflection lines of patterns made with cobalt radiation. It was not possible to make a precise determination of the lattice parameters of Zr2Au because the filings underwent rapid oxidation at room temperature. The evidence supporting the large tetragonal cell containing 6 atoms involves the presence of the reflections 002, 101, 004, 112, 114, 202, and other reflections with higher indices for which I = 3n. Three or more reflections corresponding to these indices, which cannot be indexed on the basis of a small body-centered tetragonal cell containing 2 atoms, are observed in the patterns of all of the compounds listed in Table I. They are strongest and most numerous when the components have large differences in their atomic scattering factors, as is the case in Ti2Pd, Zr2Cu, and Hf2Cu. The values of the z parameter should be regarded as approximate They were selected on the basis of the best agreement between the observed intensities of two pairs of closely spaced powder pattern lines (004): (112) and (114): (202) and the corresponding intensities calculated for several values of z. An ordering scheme was assumed in which titanium, zirconium, or hafnium atoms occupy the titanium positions of Ti2Cu, and copper, silver, palladium, or gold atoms occupy the copper positions. Calculated and observed intensities of the aforementioned lines and the other lines of the patterns are in satisfactory agreement, but a refinement of the z parameters and a final confirmation of the ordering arrangement must await a more detailed study. It is interesting to note that in transition metal systems at A2B stoichiometry the Si2Mo-type structure competes for stability with two other fre-

Jan 1, 1962

By D. H. Baldwin, R. E. Reed-Hill

DURING the course of experiments involving oxygen equilibrations with a high-purity Pd-5 at. pct Rh alloy, the appearance of a subscale was noted. Most of the heat treatments in a pure oxygen atmosphere were performed at 900°C. The presence of the subscale was noted when the oxygen partial pressure was in excess of approximately 95 mm, up to 1 atm, the highest pressure used. According to available thermodynamic information' on the free energy of formation of Rh2O3 (the only reliably reported solid oxide of rhodium), this oxide should not be stable under these conditions, nor, of course, any oxide of palladium. Extraction of the subscale from the matrix was affected by treatment with aqua regia, and the resulting X-ray diffraction powder pattern is shown in Table I. This pattern is not that of Rh2O3, nor that of any known rhodium or palladium oxide. It can be indexed on a hexagonal lattice with a. = 5.22Å and co = 6.0Å. In addition to this unknown material, there was a weak Rh2O3 pattern. An X-ray fluorescence analysis of the powder gave 3214 palladium counts per sec and 5381 rhodium counts per sec, thus indicating the likelihood of a double oxide of palladium and rhodium. In view of the fact that the above result was obtained on a single sample, it is perhaps best not to insist upon the above ratio of about 0.6, but possibly to tentatively assume 0.5 or 1/1. The small amount of Rh2O3 in the X-ray pattern might have occurred on cooling, and it seems possible that the double oxide, below some temperature level, is not as stable as Rh2O3. To examine this point, some of the extracted double oxide was heated on a sheet of pure palladium in air at 900°C, and allowed to furnace-cool over a period of several hours. The X-ray diffraction pattern of the powder showed approximately the same intensity of Rh2O3 and the "hexagonal" oxide as before, but, in addition, a rather faint pattern of another crystalline species made its appearance. This finding again suggests partial transformation of the major (double) oxide to other oxides on cooling from 900°C. Application of the phase rule shows that only one equilibrium oxide (subscale) could have been present at temperature. Vacuum-fusion, oxygen analysis of the extracted subscale showed 7.7 ± 0.7 pct O. Combining this information with that of the X-ray fluorescence analysis suggests the formula PdRhO. This compound should have an oxygen content of 7.07 pct, which is within the range of error of the oxygen analysis. However, until a more careful stoichiometric study is made, it is suggested that the formula be written

Jan 1, 1965

By J. M. Snook, F. C. Langenberg

FOR a better understanding of the melting and refining processes, the oxygen content of the steel bath should be determined. The oxygen content influences many process and product variables such as the efficiency of alloy recoveries, the ingot structure (through reaction with carbon in the steel), certain reaction rates, cleanliness (formation of indigenous oxide inclusions), and mechanical properties. Many methods of sampling liquid steel for oxygen have been devised. These methods can be divided into two main groups: 1) taking samples directly in the furnace or ladle;'-'0 2) removing the metal from the furnace or ladle in a sampling spoon and pouring it into a mold, or drawing the metal from the spoon into a sample holder.11-l4 The advantages and disadvantages of the two general sampling methods are presented in Table I. The bomb-sampling technique described by Huff, Bailey, and Richards8 has been used by the Crucible Steel Co. in certain research projects where oxygen contents were needed. It was recognized that a localized boiling condition could result when the mold is introduced into the steel bath, although the work of Gilbert and Bailey9 seems to indicate that the carbon boil occurring in the bomb-sampler cavity has a negligible effect on the oxygen content. In order to confirm the reported negligible effect of a carbon boil, additional work was done with adaptations of the conventional bomb sampler. The methods used are shown in Fig. 1 and are as follows: 1) Type A: aluminum wire (0.125-in. diam) coil in cavity; steel cap over mouth 2) Type B: aluminum cup in cavity; aluminum cap over mouth, 3) Type C: aluminum cup in cavity; steel cap over mouth, 4) Type D: fine aluminum wire (0.013-in. diam) in cavity; steel cap over mouth. The amount of aluminum wire in the Type A and D samples varied from 0.8 to 1.2 pct of the sample weight. The aluminum cups averaged 2.0 pct of the sample weight; the latter were used to ensure that the steel contacting the walls of the bomb cavity was fully killed. Samples were taken from plain carbon heats made in open-hearth furnaces. This grade of steel before silicon additions is essentially an Fe-C alloy and the effects of other alloying elements on the oxygen activity can be neglected. The results of sampling with the bomb samplers is shown in Fig. 2 where the percent oxygen is plotted against the percent carbon on log-log paper. Spoon sample results on the same grade of steel where the sample was killed in the spoon are shown in Fig. 3. The oxygen contents were determined by the vacuum fusion method. A comparison of Figs. 2 and 3 shows that the oxygen concentrations obtained with the spoon-sampling method (metal killed in the spoon after slag removed

Jan 1, 1960

By R. A. Perry

A speiss is an artificial arsenide, sometimes an antimonide, formed in lead smelting, smelting of oxide copper ores, and in some lead-refining operations. The production of speiss is closely allied with the amount of arsenic and metallic iron in the blast-furnace charge, while dross is produced mostly from the copper and sulphur, with varying amounts of antimony, arsenic and nickel. Most speiss is essentially an iron arsenide containing some sulphur, copper and nickel, if these elements are present in the furnace. The recent introduction of the soda treatment of lead drosses has produced a speiss high in copper. Speiss is not miscible to any extent with molten copper, lead, matte or slag. Miscibility increases with temperature. This property of the speiss is clarified in the latter part of this paper. Most speisses are produced in conjunction with matte and their sulphur content points to the solubility of matte in speiss. . Some plants have insufficient speiss to warrant separate treatment. When arsenic is high in the blast-furllace charge, there may be sufficient speiss to tap separately from the first settler. Arsenic From the foregoing, it is obvious that a study of the treatment of speiss would not ,be complete without consideration of arsenic. Inasmuch as perhaps go per cent of the blast-furnace feed may be sinter, a definite knowledge of the properties of arsenical ores and compounds that go in to the roaster and sinter feeds is of value. Lcad smelters obtain their arsenic from ores and concentrates, as well as from flue dusts shipped from other plants, particularly copper smelters. The percentage of arsenic in ores is usually low, although occasional lots of high arsenical products are smelted. These crude ores may be sulphides, as arsenopyrite; or oxides like scoiodite, which carry gold or lead minerals. The small amount of arsenic in the original ores received is greatly increased by a circulating load of arsenic in secondary products at are resmelted for their lead or precious metals. More important than the arsenic present directly in the concentrates treated is that in fumes and dusts that may originate in the blast furnace or sinter plant of the lead smelter, or may come from a copper smelter as lead-bearing converter fume or roaster dust. Thus the lead smelter not only must treat its own arsenic but is compelled to handie that from some copper and perhaps other lead plants. Where copper and lead plants exchange products, each plant may be forced to re-treat part of the arsenic it has shipped ,outside its own confines. Speiss from the lead blast furnace contains chiefly iron and arsenic with some lead, copper and sulphur and often appreciable amounts of gold and silver. If its copper content is sufficient this speiss may be sent to the copper smelter where usually it is treated in the converter and the copper is separated as blister. Much of the arsenic and most of the lead are recovered as converter fumes

Jan 1, 1944

By H. L. Talbot, A. J. Hall, R. S. Young

A method for the chemical differentiation of oxidized cobalt from sulphide cobalt in complex mine and concentrator products is of considerable interest to a number of mining companies throughout the world. At Rhokana Corporation, in Northern Rhodesia, for instance, cobalt is found chiefly in the sulphide form, as carrollite, but a cobalt oxide mineral also occurs irregularly underground, and relatively small quantities in the mill feed from time to time are difficult to detect visually at the concentrator. The oxide cobalt is not amenable to the sulphide flotation process used at Rhokana and nearly all of it passes into the tailing. The occasional period when high tailing losses occur with no appreciable change in the cobalt content of the mill feed has naturally been disturbing to the concentrator staff, who had no means of determining whether an influx of cobalt oxide ore or laxity on the part of the mill flotation operators as responsible for these results. A procedure for determining oxidized copper minerals, based on their solubility in dilute sulphuric acid saturated with SO2, has been used for many years in the copper industry. In the presence of cuprite the procedure fails, but otherwise it gives results quite satisfactory for control pur- poses. Empirical methods for determining oxidized forms of zinc and lead have been proposed, but they are generally unsatisfactory except in isolated cases of a very simple ore. No published procedure for determining oxidized forms of cobalt or siinilar elements has come to the authors'notice. Experimental Procedure There are obviously two alternatives in attacking a problem of this nature. One is to employ a solvent for oxide cobalt that will not dissolve cobalt sulphide. The other is to find a substance capable of oxidizing the cobalt sulphide to sulphate, which of course is water-soluble and can then be leached from the water-insoluble oxide cobalt. A representative sample of Rhokana cobalt oxide ore was secured, ground to minus 200-mesh, and carefully rolled to furnish a standard sample for this work. The analysis of this material gave the following percentages: Co, 0.55; Cu, 1.64; Fe, 2.98; S, 0.08; SiO2, 41.42; Al2o3, 11.20; CaO, 13.82; MgO, 3.80. As a source of cobalt sulphide a similar large sample of carrollite was prepared, analyzing Co, 23.36 per cent; Cu, 27.32; and S, 31.13. This sample of course was not pure carrollite, but was the best obtainable in quantities sufficient for a standard intended for considerable work. It was a pure sulphide sample, microscopic examination showing only carrollite, a little bornite, and gangue, with no visible sign of oxidized ore present. A wide variety of reagents as used in an effort to selectively dissolve oxide cobalt or to oxidize sulphide cobalt to

Jan 1, 1949

By H. R. Hendricks

THE ore of the Iron King Mine, being very hard and having a very fine crystalline structure, presents many problems in milling that are not present in ordinary lead-zinc ores. This very fine crystallization causes much interlocking of minerals and is most prevalent between the gold and pyrite and the sphalerite and pyrite. The recovery of gold is extremely important in the economic value of the ore and therefore constitutes a large part of the metallurgical problem to be solved. All of the gold is free but it has not been observed in any of the mill products except under very high magnification. Some of the gold is tarnished, and this condition also makes recovery difficult. The gold values have been determined to be highest in association with chalchopyrite and arsenopyrite. The silver occurs mostly in association with galena and tennantite. A small amount of native silver has also been observed occasionally. The galena, being softer than the other minerals, frees much better in grinding. Some oxidation of galena is present in the ore, as the nonsulphide lead content amounts to approximately 10 pct of the total lead. Some oxidation of sphalerite is also present in the ore. Microscopic examination of the zinc concentrate shows many grains in the minus 325-mesh to be half pyrite, together with the fact that some of the zinc is marmatite, makes production of a high-grade zinc concentrate very difficult unless recovery is sacrificed. In the grinding, it has been determined that the flotation feed must be maintained at 85 pct minus 200-mesh to free the minerals sufficiently to make the best economic recoveries. Table I illustrates the effect of fine grinding. It shows that the bulk of the recovered minerals is in the minus 325-mesh size and is the part of the minerals that is more readily freed. The pyrite concentrate is not given in the table because it contains practically no minus 325-mesh. DEVELOPMENT OF PRESENT FLOWSHEET AND MILLING PRACTICE The first milling operation started in 1938, with a mill capacity of 140 tons daily, producing a bulk gold-silver-lead concentrate. After several stages of expansion and changes of the mill flowsheet, the mill is currently treating 380 tons of ore daily and three separate concentrates are produced. In order of their importance and value, these are: (1) lead concentrate, (2) zinc concentrate and (3) pyrite concentrate. The present average grade of mill heads is: gold 0.11 oz, silver 4.00 oz, lead 2.50 pct, zinc 7.60 pct, copper 0.22 pct, iron 22 pet and 30 pct insolubles. Specific gravity of the ore is 3.8. During the period when bulk flotation was practiced, the flowsheet was very simple. The ore was ground fine enough to liberate as much gold, silver and lead as possible. Necessary collectors and frothers were added in sufficient quantity

Jan 1, 1947

By Theron Wasson

The year 1942 in Michigan was devoted largely to the development of fields discovered in 1941 but was also characterized by a great increase in prospecting for new productive horizons below the Dundee limestone. Some new fields were discovered but as yet these have shown few signs of being of major importance. Michigan's total production for 1942, of 21,754,000 bbl. of oil, as reported by the U. S. Bureau of Mines, was second only to the record year of 1939 when 22,799,000 bbl. of oil were produced. All important fields were prorated throughout the year, and it is obvious that without proration a new all-time high would have been established for the state this year. From 1925 to the end of 1942, Michigan has produced annually the following amounts of oil: Year Barrels Source 1925 4,000 1926 94,000 1927 435,000 1928 592,000 1929 4,641,000 1929 Michigan Geological Survey 1930 3,928,000 3,785,000 1932 6,925,000 1933 7,942,000 1934 0,603,000 1935 15,776,000 U.S. Bur. Mines, 1937 1936 11,928,000 U.S. Bur. Mines, 1938 1937 16,628,000 U.S. Bur. Mines, 1939 1938 19,211,000 U.S. Bur. Mines, 1938 1939 22,799,000 1940 19,764,000 1941 16,361,000 1942 21,754,000 Total 183,170,000 This tabulation shows the steady growth of the industry in Michigan since the first discovery of oil in 1925. Maes A map of the oil and gas fields of Michigan in 1941, published in the report for 1942, shows the main producing areas and should be referred to if information in regard to their location is desired.* The main producing zones were shown in the columnar section in the same publication. Development in Fields Reed City Field.—During 1942 the Reed City field, which ended 1941 with 112 producing oil wells, added 70 new producers to bring the total at the end of the year to 182 wells and 4700 proven acres. A south extension of the field developed during the year was in large part responsible for the increase. Approximately 9,000,000 bbl. of oil was produced from the field during the year, bringing the total accumulated production to the end of the year to roughly 11,750,000 barrels. Headquarters Field.—The Headquarters field, which started the year with ro producing oil wells, ended the year with 41 wells aid 1240 proven acres. Approximately 2,300,000 bbl. of oil was produced in 1942. Kawkawlin Field.—Although not a 1941 discovery, the Kawkawlin field has been actively developed on the 40-acre spacing plan during the last year, 42 producers having been completed, bringing the total number of wells at the end of the year to 77. Nearly 500,000 bbl. of oil was produced during the year.

Jan 1, 1943

By D. W. Fuerstenau, L. M. Berlioz

In 1957, Charles1 first presented an analysis of a relationship between the energy expended in comminution and the extent of size reduction achieved. Subsequently, Schuhmann extended and interpreted Charles' approach to comminution. According to their analysis, the energy, E, expended in comminuting unit mass of fairly uniformly-sized material is related to the size modulus of the product, k, by where a is the distribution modulus (slope of the log-log cumulative size distribution plot) and C is a constant, which is primarily determined by properties of the solid. If there is any fundamental basis for the E vs k relationship, E must necessarily be concerned with the useful work for the size reduction process and not necessarily the work input, as discussed in detail by Harris.3 Charles and numerous other investigators tested this equation only by relating k to the grinding time, assuming expended energy to be directly proportional to time, but never directly measuring power. In our laboratories, we recently constructed a special batch ball mill with which the power could be accurately determined with torque measuring apparatus.4 With this device, a direct test of the Charles relationship could then be made for ball milling. 5 EXPERIMENTAL A stainless steel mill, 11.5 in. long and 10 in. in inside diam, lined with eight low-silhouette lifter bars, was used for these experiments. The grinding media consisted of 30 kg of single sized, 1-in. diam, stainless steel balls, which occupied about 50% of the total mill volume under static conditions. The void fraction of the static ball load was approximately 0.4, and for the dry grinding conditions used, the corresponding interstitial volume was filled when about 3300 g of the dolomite feed material was present. The 7x9 mesh dolomite used as feed was obtained as a natural, uncrushed product thus as- suring a random distribution of feed particle shapes. Charge weights of 660, 1320, 1980, 2480, 3300, 3960 and 5280 g, which correspond to 20%, 40%, 60%, 80%, l00%, 120%, and 160% of the interstitial load, respectively, were used. Grinding times of 20, 40, 60, 80, 100, 150, 200 and 300 revolutions were considered, with the mill operating at 54 rpm (approximately 60% of critical speed). The tests were made sequentially, in that the size distribution obtained after a certain grinding interval was utilized as the feed for the subsequent period of grinding. Agreement with control experiments where the mill ran continuously for the entire interval was excellent. Torque measurements were made utilizing a Baldwin-Lima-Hamilton SR-4 torque pickup in the manner that has been described previously.3 Standard wet (to obtain the -400 mesh material) and dry sieving procedures were followed using the Tyler full screen series. RESULTS The Gaudin-Schuhmann size distributions obtained in all of these experiments are characterized by distribution moduli (slopes), a, of about 0.66 and are representative of normal dry batch ball milling of a material like dolomite. As an illustration of typical

Jan 1, 1968

By Zhendong Tao, Wei Huang, Zhongxi Yang, Hang Guo, Haoran Geng, Jingxiang Yao

"CaSO3.0.5H2O typed Flue Gas Desulfurization (FGD) gypsum is a by-product formed in flue gas desulfuration by double-cycle fluid bed method, its composition is different remarkably from that of gypsum yield by normal desulfuration methods, and few studies using it as cement retarder have been carried out. In this paper, the experimental study using CaSO3•0.5H2O typed FGD gypsum as cement retarder was carried out. The experimental results indicate that, when FGD gypsum is added solely into cement, the retarding action is not satisfactory because of its slow dissolution rate; The composite retarder composed of FGD gypsum and natural gypsum can regulate efficiently setting time of cement over a wide range of addition and rate of both each other, and the mechanical property of cement can be improved significantly.Fire-power plants have discharged a lot of harmful gases such as SO2 etc., which has resulted in serious pollution on the environment. To reduce harmful gas containing sulfur in the flue gas, various flue gas desulfuration (FGD) techniques have been developed. At present, the FGD method used widely is to add desulfuration agents such as limestone powder or other calcareous-contained materials during the coal burning or in the flue gas so that the components containing sulfur react with them to form solid sulfate or sulfite and to be separated from the flue gas by collectors. CaSO3•0.5H2O and CaSO4•2H2O typed FGDgypsum can be obtained respectively by dry and wet FGD process[1,2]. As a by-product of fire-power plant, features of of FGD gypsum and its applications have been paid great attention. Pavel Tesa'rek et al [3] presented in full the characters and properties of FGD gypsum; Zofia Konik et al[4] prepared expansive additive used to cemnent industry by using FGD gypsum and aluminate clinker; A. K. Alva et al [5] investigated the effects of FGD gypsum and fly ash on soil properties and found the growth of seedlings of either rootstock improved significantly in soils amended with FGD gypsum and fly ash. FGD gypsum also can be used in gypsum binders, plasters and plasterboard manufacture [2.6]. C. S. Poon et al [7] investigated the effects of FGD sludge on the pozzolanic reaction of reject-fly-ash-blended cement pastes. Many studies using CaSO4•2H2O typed FGD gypsum as cement retarder has been reported[2,8-12]. Howevere, few studies using FGDG as retarder were reported so far. In present study, the chemical compositon, phase composition, structure and features of CaSO3•0.5H2O typed FGD gypsum have been analyzed, and its application characters used as cement retarder have been investigated experimentally."

Jan 1, 2008

By J. W. Borough

THE electrical conductivity of pure copper is markedly decreased by small amounts of impurity in solution. The magnitude of this effect has been very carefully determined by numerous investigators who studied a wide variety of impurity elements. Much of this work is summarized in an excellent review by Pawlek and Reichel.' The case of zinc impurity in copper is an exception to the above in that there has been no well-documented study of the effect using high-purity starting materials. The values for the effect of zinc on conductivity, which are usually taken from unpublished work by Skowronki, indicate that zinc in low concentration has no effect whatever in lowering the conductivity of pure copper. This effect, if true, is unusual since it is generally known that even in lowest concentration, impurities exert an effect on conductivity which is proportional to the amount dissolved. It is more reasonable to suspect that in the original work on the effect of zinc some of the zinc was oxidized out of solution during specimen preparation. The following experi- and calculated d-spacings based on the unit cell dimensions of Baenziger et al. Since the single crystal studies of Baenziger et al. were limited to (hk0) and (hk 1) data, it was not possible to Compare all the intensities of the powder and single crystal patterns. However, the strongest reflections of both studies, the (321) reflection, agree very well when the appropriate multiplicity factor is used. Good agreement was also found in a number of other reflections of lesser intensity. Despite the precautions taken in annealing the U6Fe powder, an appreciable amount of UO, was detected in the powder pattern after heat treatment. The powder pattern reported by Groganoand Clews is given in Table I. Except for the 3.15A line, which is the strongest line of UO,, their pattern, as far as it goes, compares favoEably with the data from this study. Gordon and Kaufmann have reported a diffraction powder pattern purported to be U, Fe but it bears very little resemblance to the patterns of Table I. ments were carried out to redetermine the effect of zinc on conductivity using high-purity starting materials and processing conditions capable of producing oxygen-free alloys. Materials used in the experiment were ASR 99.999 pct Cu and Cumingo 99.999 pct Zn. Four 1 lb melts were prepared with identical procedure in argon atmosphere with crucibles made from Speer No. 580 graphite. The compositions were 0.0, 0.1, 0.3 and 1.0 pct Zn. Standard No. 14 wire (0.064 in. diam) was made from these melts as follows. The 1 pct Zn sample was first reduced 94 pct by cold swaging with five annealing treatments of 800o for 10 min in hydrogen between swagings, and then reduced 94 pct to the

Jan 1, 1962