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By C. W. Thompson

The two-way belt conveyor for coal mine service simultaneously carries coal from faces and transports men and supplies into the mine, largely eliminating the necessity for rubber-tired supply and man haulage equipment and the necessity of maintaining additional headings. SINCE 1920 use of belt conveyor transportation has increased steadily, and capacity in tons per hour per unit has increased at about the same ratio. Belt conveyor systems for simultaneous two-way material transportation have been widely discussed, but heretofore have proved impractical for specific operations. Studies of a proposed means of disposing of wash-ery waste on the return trip of a main haulage conveyor, as yet uninstalled, led to the application of two-way belt haulage to Weirton's underground supply and mantrip problems. Except for a single-tracked main haulage entry, Weirton mine is essentially a full-belt haulage operation with attendant problems of supply and man-trip haulage that could not be handled with conventional belt conveyors during the regular work shift. It was necessary, therefore, to maintain rubber-tired haulage equipment and haulageways serving multiple-shift continuous operations. The conventional underground belt conveyor is primarily intended to transport coal from working faces to surface. Belts are usually reversed to haul men to face areas at the start of the working shift or to transport supplies. The two-way belt conveyor for coal mine service is not yet offered for sale as a unit by any manufacturer and must be built on order. This conveyor, with return run traveling parallel but opposite in direction to the outgoing or coal haulage run, permits not only haulage of coal from faces, but simultaneously permits transportation of men and supplies into the mine at any time, see Figs. 1 and 2. This eliminates, to a great extent, the necessity for rubber-tired supply and man haulage equipment and the maintenance of additional headings. Unlike conventional installations the return belt is in the open for inspection and maintenance. It is less subject to edge wear, since it runs on regular troughing idlers, see Fig. 3. The only points where the edge of the belt can rub are at the drive unit and the tail units where improved design should eventually eliminate edge wear entirely. Individual conveyor sections are lighter, since all that is needed is a tubular stringer on each side to carry troughing idlers only, no cover plates being necessary. The return run is carried on an assembly identical to that carrying the coal haulage run. These assemblies are definitely easier to handle and are cheaper than installations for the same length of conventional belt conveyor. The four pipes required for the complete conveyor frame may serve only as supports, or, with proper couplings, may be used for mine water discharge line, pressure line for face spraying, telephone cable conduit, or any other purpose requiring a pipe line, see Fig. 4. Additional expense, if any, for the increased number of pulleys needed to turn the belt over for the return service is more than offset by reduced expense for rubber-tired supply and mantrip haulage. The two-way belt haulage system is confined to a single heading on intake air, isolated from all other headings by stoppings on each side, and open for air circulation only at the main haulage way and at the faces. One additional heading, also on intake air, is maintained as an emergency rubber-tired haulage way, all other headings being center-posted and used as airways only until retreat work begins. This method, of course, can and should be used with any belt conveyor, since it provides two independent intake airways for escape in case of fire on the belt or elsewhere. Belt heading is heavily timbered even under good top conditions and space between the two runs permits center posting for the full length

Jan 1, 1954

By W. P. Lass

Discussion of the paper of W. P. LASS, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 103, July, 1915, pp. 1443 to 1447. ROBERT M. KEENEY, Somersville, Conn. (communication to the Secretary *).-There is not only a possibility of improvement in present methods of treatment of the byproducts of refining in the cyanide plant by the use of the electric furnace, but also in melting and refining of gold precipitate. This interesting paper by Mr. Lass shows an electric furnace in successful operation in a large plant for the treatment of by-products. Electric melting of cyanide precipitate is a cheaper form of treatment in many cases than any of the combustion methods now used because of the high cost of oil, coke or coal in many camps, while on the other hand electric power may be had cheaply, especially in a place where there is possibility of development of high-head hydro-electric plants for power purposes. Also, if the oil-fired tilting furnace is used for direct melting of the precipitate, as is the case at many small mills, a great saving in cost is made by the use of the electric furnace because of the high crucible cost in melting with the tilting furnace. A study of this paper indicates the cost of the operation approximately. Consider all the fluxes to be turned in as byproducts and coke at $18 per ton.

Jan 12, 1915

By D. G. Hartley, P. C. Hayward, K. R. Weller, U. J. Sterns

The major copper loss at Mount Lyell is in the coarsest particles presented to flotation. An analysis of data collected from the five Primary and six secondary ball mill circuits showed that grinding was limited by low breakage rates in secondary mills, possibly due to oversize balk. For five months, the make-up ball size for one secondary mill was reduced from 50 to 30 mm (2 to 1.1 in.). Frequent steady-state tests indicated that overall average breakage rates could be increased by about 70% using smaller balk. A mathematical simulation of the whole circuit with increased breakage rates showed that the resultant finer grind was likely to improve copper recoveries by about 2%.

Jan 1, 1984

By T. W. Healy, D. R. Dixon

The mechanism of activation of ZnS by metal ions and the subsequent flotation induced by xanthate has been considered as a surface redox reaction. By ultraviolet (UV) irradiation of ZnS suspensions, the normally inert ZnS is caused to react with xanthate in the absence of activator. By similar UV irradiation, the rate of adsorption-exchange of activator ions at the ZnS-H2O interface is also increased. Mechanisms of these reactions are proposed and compared to current electrochemical and semiconducting models of sulfide flotation.

Jan 1, 1976

By M. H. Caron

D. C. Ralston—The fact that none of the organizations that have worked on these ammoniacal leaching processes have contributed discussion of Mr. Caron's papers today is a matter of some disappointment. Adaptations, modifications and improvements have been made by several organizations or individuals and to complete the record I want to document a number of them. The Nicaro Nickel Co. operated a Government owned plant at Nicaro, Oriente, Cuba, during the past war. A report on the project was written for Reconstruction Finance Corp. and is in the document file collected by the Office of Technical Services, Commerce Department and filed with the Library of Congress, where it can be consulted or bibliofilm or photostat copies obtained by asking for copies of PB 97271. Nicaro Nickel Co. has obtained a series of U.S. Patents, my list (probably incomplete) of which bears the following numbers: 2,400,098; 2,400,114; 2,400,115; 2,400,461; 2,400,612; 2,458,902 and 2,473,795. International Nickel Co. has also studied the problem and U.S. Patent 2,478,942 shows that some of the complications involving occasional low extractions have been unraveled. Forward, Samis, and Kudryk, of the University of British Columbia, have also published a study of copper-nickel sulphide concentrate.' Others are interested and it is evident that this type of hydrometallurgy is likely to find a happy application in the not-too-far distant future. Caron's two papers practically amount to a summation of his life's work on the important subject of nickeliferous iron ores and the American Institute of Mining and Metallurgical Engineers is fortunate to have been chosen the medium of publication. M. Rey—The ammonia leaching process has been the subject of certain studies in New Caledonia, but at the present time, the conditions are not very favorable to its adoption. The process is a combination of a thermal and a hydrometallurgical process. As a result, the first costs and the operative costs are rather high. To counteract this unfavorable situation, it would be necessary to make a very high extraction. But unfortunately, the process is delicate and in spite of all of Caron's valuable work, there are still many unknown factors. It is difficult to obtain regularly a high extraction, exceeding 75 to 80 pct. It would be interesting to have the opinion of Inter-national Nickel on the process and to know what the Dutch are doing in Celebes. M. H. Caron (author's reply)—I am quite familiar with the different types of nickel and cobalt ores as found on New Caledonia, and it may be pointed out that the Celebes ores are quite the same in their properties. Their behavior when treating the ores by the ammonia-leaching process is also alike and in my opinion, there is no reason why a high extraction of the New Caledonian ores could not be obtained if the process is properly applied. For the sake of completeness I may add that one of my patents has not been discussed in the papers. This patent deals particularly with the treatment of serpentine nickel ores. (Netherlands No. 62693 40a, French No. 965.786 April 21. 1948. and Cuban No. 13.625 Jan. 1950.)

Jan 1, 1951

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons introduced by members.) Mining Engineer. Graduate of Colorado School of Mines, 1912, experienced in coal, lead manganese and limestone mining, wishes execu¬tive position. A-2146. Graduate Mining Engineer.-Discharged, reserve officer, age 32, single, 7 years' experience holding responsible executive positions, in iron, copper, and talc mining. Well recommended. A-4913. Mine Manager, Superintendent or Engineer. Age 41, graduate of Michigan College of Mines; 20 years' experience in responsible positions in connection with large operations; in iron, copper, pyrites, lead and zinc mines; successful in finding and developing orebodies, increasing production, and reducing costs. Desires a position with a future. A-4081. Mining Engineer.-Graduate from Polytechnicum, Stockholm; 14 years' practical experience in Sweden, United States, Belgium, and Spain, in iron-ore development, mining, and milling; thorough knowledge of English, French and Spanish; desires position as superintendent or assistant manager. Good references; member; now in Sweden. A-4914. Research Engineer. Now in Chile on the erection of a 50-ton mill with modern machinery for the treatment of Chilean nitrate. Desires connection from the first of next year; with firm operating in South or Central America. A-4915. Metallurgical Engineer. Junior member, age 26, single, graduate Columbia School of Mines; recently discharged as lieutenant in the heavy artillery. A-4916. General Superintendent. Technical graduate, age 35, desires position over a group of four or five mines or as assistant general superintendent for a large. company. Has had experience in superintending mines, construction work, opening new and abandoned mines, safety and effi-ciency engineering and handling organized labor. At present employed by a coal company in the middle West. A-4917.

Jan 11, 1919

By T. M. Drown, P. W. Shimer

The precipitation of phosphoric with titanic acid, by boiling an iron solution which had been reduced to the ferrous condition by sulpharetted hydrogen or sulphurous acid, was first noticed by E. H. Bogardus in 1874.* Since that time, as far as we are aware, not much has been published on the relation of these two acids to each other, and to silicic acid, in the ordinary course of analysis of iron ores. The following investigation may perhaps aid in clewing up some of the obscure points in the analysis of titaniferous ores. THE DETERMINATION OF PHOSPHORUS. From two to five grams of the finely powdered ore are weighed into a beaker and treated with about 50 c.c. of hydrochloric acid (sp. gr. 1.12), evaporated to dryness, and heated ih an air-bath for

Jan 1, 1882

By R. A. Burmeister, G. P. Pighini, P. E. Greene

An open tube vapor phase epitaxial growth system has been used for large area (multiple substrate) growth of GaAs1-xPx on GaAs substrates. The GaCl-GaCl transport reaction is used with either a GaAs or Ga (nonsaturated) source. Selenium and tellurium have been used for donor impurities, and zinc as an acceptor. The useable substrate area in this system is approximately 20 sq cm. The uniformity of thick-ness of the epitaxial layers are typically better than ±5 pct across a given wafer. Electrical and optical measurerments indicute comparable uniformity in electrical and luminescent properties within a wufer. The application of this system to the large scale pro-duction of GaAs1-x Px for display devices, both discrete and arrays, is discussed. Typical electrical and luminescent properties of light emitting diodes fabricated front material produced by this technique are presented. THE most promising materials currently being utilized for visible injection electroluminescence are GaAs1-xPx, Ga1-xAlxAs, and Gap. All have reasonably efficient emissions in the red portion of the visible spectrum at room temperature; Gap also has an efficient green emission.' At present, GaAs1-xPx has a technological advantage over Ga1-xAlxAs and Gap for display applications, since relatively large (several sq cm) areas of GaAs1-xPx suitable for use in electroluminescent devices may be readily grown by vapor phase growth techniques. In contrast, the preparation of Gap and Ga1-xAlxAs for electroluminescent device applications generally employs solution growth techniques which are at present not well suited for the growth of large areas of uniform thickness and doping level. The capability of uniform growth over large substrate areas and the use of multiple substrates is necessary for the practical utilization of electroluminescent devices. This is particularly important when quantity production or monolithic devices are required. Furthermore, in many display applications arrays of light emitting devices are used, the individual elements of which are of a size resolvable by the unaided eye. Thus the overall dimensions of display are substantially larger than those of most semiconductor devices. It is also necessary to achieve a high degree of control over the growth parameters to attain the required degree of reproducibility of materials properties for electroluminescent devices. In the case of GaAs1-xPx it is necessary to accurately and precisely control the phosphorus content of the alloy, both on a macroscopic and microscopic scale, in addition to the parameters generally associated with epitaxial growth such as thickness and doping level. This value is critical, as it has a major effect on the performance of electroluminescent devices. This paper describes the epitaxial growth of GaAsl-xPx suitable for electroluminescent display devices using a system developed specifically for this purpose, and which contains several novel features. The results of studies of selected physical properties of the epitaxial layers are also discussed. Finally, a brief summary is given of the characteristics of display devices fabricated from GaAsl-xPx grown in this system. EXPERIMENTAL A) Reactants. A number of techniques suitable for the vapor phase epitaxial growth of GaAs1-xPx have been reported in the literature.'-' The method selected for this investigation is that in which the Ga is transported by the GaC1-GaCI3 reaction in an open tube process. The results reported here were obtained using either the combination of GaAs, AsC13, and pH3, or Ga, AsH3, pH3, and HC1 as the initial re-actants.* The ASH3 and pH3 were obtained as dilute *Several different sources of supply were used for these reactants, y~elding comparable results._____________________________________________________ mixtures in HZ; the HC1 was obtained from the reduction of AsC13 by Hz at elevated temperatures. Both selenium and tellurium were employed as donor impurities, and zinc as an acceptor impurity. Selenium was introduced in the form of H2Se, tellurium in the form of tellurium-doped GaAs, and zinc in the form of diethy1 zinc. B) Apparatus. The prinicipal difference between the apparatus used in the present study and that of Tietjen and Amick,8 in addition to size and other related design features, is that RE induction heating is utilized in place of resistance heated furnaces. Induction heating was selected for this application because it appears to have several advantages, including: 1) It is possible to keep all fused silica portions of the apparatus at temperatures well below those of the reaction zone, thus minimizing a possible source of contamination. 2) The thermal mass of an induction heated system can be made small, thus reducing the total time required for the growth process. 3) Sharp temperature profiles (desirable for high deposition efficiency) are easily achieved. 4) The volume of the system for a given substrate area can generally be made smaller than a comparable resistance heated unit. This results in shorter time

Jan 1, 1970

By L. Douglass Anderson

Fox some years past the annual reviews of the metallurgy of lead have almost uniformly stated that there have been no great changes, such as there were, being more particularly noticeable in the refinement of details. While this is quite true, nevertheless these changes in details are gradually bringing about a condition that, taken as a whole, constitutes a considerable advance in the art. Local circumstances have often -compelled metallurgists to adopt certain peculiarities of practice which -upon trial have been found advantageously applicable to more than the restricted district f origin. This paper will deal ' with some of the more novel features developed in recent years at the United States smelter, Midvale, Utah, particularly with reference to the influence of the bag house. The plant is equipped with six blast furnaces, 48 by 160 in., mechanically fed; four Wedge furnaces for roasting matte and sulphide ores; four Dwight-Lloyd sintering machines; 20 converter boxes for blast roasting; storage bins for ore, fuel, and fluxes; "oxide" and "sulphide" sampling mills; bag house taking all the gases of the plant; arsenic plant for treating the dust gathered there; laboratory, shops, etc. The most marked feature of the plant is its complete bag-house system. Of the many schemes thus far tried for preventing damage to surrounding agricultural interests the writer is aware of none as yet quite as efficient as the bag house. A common remark of visitors upon approaching the plant and viewing the clear stacks is that it appears to be shut down, whereas once inside all the evidences of industry are found.

Jan 7, 1914

By Henry H. Angst

Tim history and efficiency of 40-ton capacity dump trucks for surface waste removal at the New Cornelia opencut copper mine, at Ajo, Ariz., are summed up in this paper. Tabulations of truck performance, together with detailed cost figures, prove the value of truck transportation for certain special operations at a large mine. Emphasis is given to the fact that truck transportation not only reduces the haulage costs but also reduces the breaking ground and shovel-loading costs per ton. The New Cornelia opencut copper mine is in the desert. The climate is ideal for seven months of the year, but the five summer months are uncomfortably hot. The elevation is 1820 ft. above sea level and the annual rainfall is approximately 6 inches. The ore body is a monzonite-porphyry deposit. The waste-rock capping consists of hard rhyolite, conglomerate and porphyry. All the material is hard and cannot be excavated without blasting. The rhyolite and conglomerate waste is very blocky. The weight of the rock in place is 2.16 tons per cubic yard. The average mine production per day is 22,000 tons of ore and 18,000 tons of waste rock, the latter consisting of surface rock and material within the ore body too low in grade to send to the concentrator. The rock is loaded by electric shovels with 4 1/2-cu. yd. dippers. The primary transportation of ore to the concentrator and waste to the dumps is by 30-yd. dump cars and oil-burning steam locomotives; therefore truck haulage of waste is a minor factor of the New Cornelia operation. However, as New Cornelia was one of the first mines to use large-capacity dump trucks for haulage, the results obtained are of interest. EQUIPMENT On the south side of the New Cornelia is a hill rising about 500 ft. above the normal surface elevation, roughly 1000 ft. in diameter at its base, called locally Arkansas Mountain peak. Exploratory drilling in this area developed additional ore that could be made accessible for open pit operation only by moving the mountain face a maximum distance of 400 ft. to the south. The slope of the mountain adjacent to the pit averaged 35°, and the small area involved precluded the use of train haulage for the removal of the 4,000,000 tons of waste estimated above the normal pit elevation. A study of dump-truck haulage at Boulder Dam and San Gabriel Dam led to the decision to use large-capacity end-dump trucks rather than smaller trucks, for the following reasons: I. The truck would have to withstand the impact of hard, heavy chunks, some of them weighing as much as 10 tons, which would be dumped from the 4 ½ -cu, yd. dipper. 2. The coarse, rocky, abrasive material would soon damage a light truck body. 3. The body of a small truck could not be efficiently loaded with coarse rock.

Jan 1, 1941

By Michael Tenebaum, C. C. Brown

In recent years, the importance of slag control in basic open-hearth operations has been universally recognized. To effect such control during the working period of the heat, methods have been developed to measure or estimate slag oxidation and slag basicity. While satisfactory rapid chemical methods have been developed to determine slag oxidation, no completely adequate method of evaluating basicity has yet been provided. The most generally accepted control methods of measuring slag basicity are based on visual cstimates. While considerable benefit may be derived from such estimates, the need for some measurable indication of slag basicity has become increasingly apparent. In a recent paper, Philbrook and Jolly suggested a promising technique for making the basicity determination. Their method consists of measuring the pH of a mixture of powdered open-hearth slag and distilled water. It was found that a relation could be established between the slag basicity and the alkalinity of the water mixture as measured by the hydrogen-ion concentration. The general procedure followed in making the measurements and the considerations involved in constructing the original curves are given in the paper of Philbrook and Jolly and need not be repeated. It is the purpose of this paper to discuss some of the factors considered in applying the pH measurements to actual slag control and to present some of the results of its apphcation. Factors Involved in Measuring pH of Water-slag Mixtures To facilitate the application of the pH control system, it was desirable that no alteration be made in the routine methods of sampling and crushing the furnace slag. Some of the factors considered before establishing a curve to relate slag basicity with the pH reading arc reviewed individually in the following paragraphs. Effect of Cooling Rate Under actual operating conditions there can be considerable variation in the temperature of the piece of slag crushed for the pH determination. To determine whether variations in crushing temperature or in cooling rate would affect the pH reading, a series of four samples was prepared from each of two separate slags. These samples represented conditions far more extreme than would actually be encountered. To obtain these samples, a thick pancake of slag was poured. One piece of this pancake was crushed immediately after it had solidified. Another piece was water-quenched and then crushed. A third piece was allowed to air-cool to room temperature before crushing. The fourth sample was prepared from a chilled slag. This sample was poured onto a heavy inclined steel plate at the

Jan 1, 1945

By W. D. Klopp, R. I. Jaffee, H. R. Ogden, D. J. Maykuth

The purification of commercial-purity tantalum powder by vacuum sintering in the temperature range 2600° to 2860°C has been investigated. Mixtures of tantalum oxide and tantalum carbide were sintered at 1800° and 2200°C. Tantalum bars containing 1 pct additions of oxygen, nitrogen, and carbon. were sintered in the range 1800° to 2600°C. Anulyses were conducted on the sititeyed mixtures to determine the extent of removal of oxygen, nitrogen, and carbon at each sintering condition. The mechanism for remova1 of each impurity, and the conditions necessary for preparing ductile tantalum by vacuum sintering are discussed. INTEREST in the refractory metals has been stimulated in recent years by the need for structural materials capable of operating at temperatures of 1100°C and higher. These metals include columbium (niobium) (mp 2468°C), molybdenum (mp 2620°C), tantalum (mp 2996°C), rhenium (mp 3160°C), and tung- sten (mp 3415°C). In addition to its high melting point, tantalum possesses the advantages of forming a solid, nonvolatile oxide at temperatures of interest and of being ductile and easy to fabricate. The classic method of consolidating tantalum from powder refined from the ores is by vacuum sintering at temperatures on the order of 2600°C. This sintering treatment both densifies the compacted powder bars and purifies the metal by causing impurities to boil out. The present work was undertaken to provide more quantative data on the relationship between sintering conditions and purity and to determine the mechanisms of impurity removal.

Jan 1, 1961

By S. L. Crouch, C. Fairhurst

The general term "coal mine bump" refers to the sudden and violent failure of in-situ coal. Coal bumps occur in most countries where coal is worked by underground methods. They are related to geological factors and mining techniques. Coal bumps can be characterized as unstable releases of energy associated with yielding that occurs with progressive mining. Unstable releases of energy occur when the coal is able to absorb less energy than is released by the surrounding rock mass during the yielding process. At any stage of mining, the amount of excess energy so released is governed by the mining geometry, the overburden stress, the elastic properties of the rock mass, and the stress-strain characteristics of the seam material. If the coal in the vicinity of the working face is crushed, the peak abutment pressure occurs at some distance into the intact portions of the seam, and mining takes place in essentially destressed coal. With subsequent mining, additional yielding occurs, excess energy may be released, and the peak stress is again shifted into the solid. This process continues until the mining geometry is such that the zone of peak stress is effectively surrounded by yielding coal, at which point it can be anticipated that further mining will produce dramatic redistributions of stress and be accom¬panied by large releases of energy. On this basis, the differences between several alternatives for mining a given block of coal may be pronounced.

Jan 1, 1975

By J. S. Ross

OIL-FIELD waters enter into many underground problems with which the petroleum engineer has to deal. Whether the problem is one of infiltration or natural encroachment, it is always desirable to determine the source of the water before commencing repair or remedial work. The usual mechanical tests for determining the source of extraneous water in wells are slow and expensive. An excellent means for the identification of oil-field waters may be afforded by their chemical characteristics when the chemical character of waters from the various sands has been established, by careful sampling and analysis cautiously interpreted. In the Salt Creek field, Wyoming, a fairly high degree of efficiency has been attained in the controlof field operations by utilizing the information afforded by the analysis. All of the operating companies appreciate this means of identifying waters, with the result that water-sampling has become general practice. Most of the chemical work dealing with Salt Creek waters has been done since the laboratories of the Midwest Refining Co. and the U. S. Geological Survey, were established in the field; the former in 1922, the latter in 1924. A spirit of cooperation has existed between the company and the government personnel by which many of the data for this paper were made available.

Jan 1, 1928

By Raymond E. Riethmann, Beuford M. Bunnell

The spiral classifier was originally developed for closed circuit grinding It has since been applied very successfully to other classification duties where a two-product size split is required. Inherent operational qualities of this classifier make it ideally suited for sizing, washing, or dewatering materials such as ores, phosphate rock, sands, coal, oyster shells, solar and potash salts. Operating characteristics, flexible design and guidelines for proper application are presented.

Jan 1, 1978

By Walter Mueller

Transition in the underground coal industry, begun with the passage of the Federal Coal Mine Health and Safety Act of 1969, should be completed midway in the 1972-80 period. Surprisingly, many of the far reaching effects of this act have been, for the most part, beneficial to the industry from the standpoint of health, safety and basic operating standards. Production has dwindled, but the next few years will see production declines level out. Average declines of approximately 27% have been recorded since passage of the 1969 act, with costs increasing an average of $1.50 per ton throughout the industry. Nationwide averages presently are in the range of 14 tons per man-shift, and by 1980 this may be further reduced to a 10-ton per man-shift average for the industry. But in spite of the many temporary set-backs affecting the industry, the marketing opportunities are clearly visible. Due primarily to the energy squeeze facing the U.S. over the near term, the 'underground coal operator is and will be enjoying unusual leverage in selling his product.

Jan 10, 1972

By Claudia Gasparrini

Silver occurs in nature in major, minor and variable amounts in a very large variety of minerals and in several mineral species in the same ore. Because these minerals may respond differently to the same flotation or leaching agents, detailed studies to establish the correct silver mineralogy become essential for optimum recoveries in those situations where the silver is treated as a primary product. On the other hand, fine grain sizes, which often affect gold recoveries, are not generally as critical in silver metallurgy, due to the lower market price which allows for more of the element to remain In the tailings before incurring in economical losses.

Jan 1, 1984

By S. Zador, C. B. Alcock, K. Sudo

Measurements have been made of the standard free enevgies of formation of the sulfates of cobalt, Cupric copper, nickel, and ferric iron in the temperature range of 750° to 950°C. Dynamic gas-solid techniques have been employed in which oxide-sulfate and basic sulfate mixtures were placed in a stream of gas of controlled SO3 potentials. Thermal analysis and thermogvavimetry were used as detection methods, and it was found THE formation of sulfates during the roasting of sulfide ores is now considered to arise in many cases from the interaction of sulfur trioxide in the gas and the metal oxide which is formed on the sulfide as a first step.' As a preliminary to the investigation of the kinetics of sulfate formation during roasting, measurements have been made of equilibria from which the free energies of formation of cobalt, copper, nickel, and iron sulfates can be that the range of temperatures which could be investigated and the choice of method was largely controlled by the kinetics of these gas-solid reactions. The results are compared with previous results obtained from static measurements with which they agree quite well, and calculations have been made employing themal data to interrelate the present results with those for other sulfates. calculated. The method used has been to expose the solid sulfate (or basic sulfate) to a controlled flowing gas mixture of SO,, 02, and SOs and to observe the temperature of decomposition. The onset of decomposition was detected by thermal analysis or thermogravimetry. EXPERIMENTAL METHODS The thermal-analysis method was similar to that employed by Dewing and Richardson.' The sulfate was dehydrated at 250°C in air, and placed in a small alumina crucible which was held by a quartz ring in the hot zone of a horizontal furnace. The reaction tube of this furnace was made of mullite.

Jan 1, 1965

By H. R. Hose

Total crude fluorspar reserves in western Europe, containing more than 35 pct CaF*, are estimated at 12 million metric tons, while reserves in the USSR and Soviet sphere may exceed 5 million metric tons. The bulk of these reserves are concentrated in the USSR, Spain, United Kingdom, Western Germany, and Italy. Since published estimates of the tonnages of individual mines and explored deposits are not available for most countries in Europe, the estimates given in Table I for individual countries are necessarily approximate, as they have been compiled from various sources and modified and combined according to the author's judgment. The majority of fluorspar deposits in Europe belong to the late Paleozoic metallogenetic epochs and have been found within or along the borders of the Hercynian massifs. Deposits in the Central and Esterel massifs of France, those of Western and Eastern Germany, Sardinia, the Catalonian coastal range and the Sierra de 10s Santos in Spain, and those of Devonshire and Cornwall belong to the Hercynian metallogenetic epochs. The deposits in southern Norway and Sweden belong to the middle Paleozoic or later, those of northern England to the late Paleozoic or later, those of northern Spain and Greece to the late Mesozoic, and those of northern and probably central Italy and Switzerland to the early Tertiary. Fluorspar occurs typically in veins, lenses, and stringers, but also in bedded replacement deposits and pockets, and may be colorless to white, yellow, green, pink, brown, blue, or purple to almost black. Except for the colorless variety, the only kind which can be used for the optical industry, the color of the fluorspar does not affect its ultimate use,

Jan 4, 1955

By C. P. Parsons

CALIPER logging is a practice of measuring the variations in the diameter of the open hole in a well. This information is useful for many purposes, among which are: Determining the volumetric capacity of the portion of open hole to be cemented; the amount of gravel to be used for gravel-packing a well; the size and location of seat for a packer to be placed in a well; studying the effects of nitroglycerin shots; studying the effects of acidizing; studying the cause of drill-pipe failures; studying drilling technique; geological correlations. This is a new practice in oil fields and is eliminating considerable guesswork.

Jan 1, 1942