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By T. Wasson

The same limitations' that characterized 1931 existed in the old fields of Indiana and Illinois during 1932. Prices remained low. The average for the year for Illinois and southwestern Indiana was $1.032 as against $0.852 for 1931 and $1.616 for 1930. In September a curtailment of production of about 25 per cent became effective and was continued to the end of the year. There were no new discoveries and very little wildcatting was done. Considering the age of the wells, and the number still producing, there were few abandonments. The most outstanding thing about the old fields in these states is their long continued productivity. Illinois The fields along the southeast end of the LaSalle anticline had their peak of 33,686,000 bbl. in 1908. In 1932, with a 25 per cent restriction for 355 months, these fields produced 4,661,000 bb1.l There were 52 new completions (most of them in old territory) and of these 33 were dry, 8 were gas, and 11 were oil wells. The new oil wells had a total initial of 102 bbl., or an average of 9.3 bbl. per well. Nineteen of the new producers were in Crawford County. There were about 200 wells abandoned in the old southeast fields in 1932 as compared to 210 in 1931. The greatest number, 86, were in Lawrence County. To the end of 1932, the four major counties—Clark, Crawford, Lawrence and Wabash—had produced almost 394,000,000 bbl. from 92,160 acres. Better than one-half of this was produced in Lawrence County from 26,240 acres. The Dupo area of Sinclair County had no additions to production although four wells were drilled. On the Jacksonville structure in Morgan County, six gas wells with an average depth of 300 ft. were drilled to sands in the base of the Pennsylvanian. They produced an open-flow total initial of 2,000,000 cu. ft. This gas will be developed further to supply a local market.

Jan 1, 1933

By P. Bandopadhyay, T. C. Rao

Geologically the Indian coals belong to Gondwana formations. Being of drift origin, these coals contain high percentage of mineral matter distributed in finely disseminated form. Indian coal-preparation plants process prime coking and medium coking coals for use in steel plants. Limited resources of good quality coking coals (Table 1) could not keep pace with the rapid developments of the iron and steel industry in India. This resulted in the introduction of coal-preparation technology and the establishment of washeries in India in the last 30 years or so. The first four washeries were set up during the early fifties at West-Bokaro, Jamadoba, Lodna and Kargali. Next came the first non-coking coal washery at Nowrozabad set up by the Associated Cement Company. In 1961, following the recommendation of Central Coal Washeries Committee (1953-54), Hindustan Steel Ltd. had set up five central washeries at appropriate locations, viz Durgapur, Dugda (2 washeries), Patherdih and Bhojudih, so that the raw coal feed having similar washability characteristics can conveniently come from a number of collieries to a particular washery. Subsequently, to meet the increasing demand of metallurgical coal by steel plants, two more central washeries (at Durgapur and Chasnala) and three pit-head washeries (at Kathara, Swang and Gidi) were installed by 1970. The prime objective of washing Indian coking coal is essentially to reduce the ash content. All the washeries in India, taken together have an installed capacity of about 26 million tons of raw coal input per annum. Details of these washeries in respect of their annual capacity, main washing systems adopted and the nature and size of the products are given in Table 2. It can be observed from this table that (1) bulk of the coal is processed in units like jigs HM baths and HM cyclones; (2) only two of the washeries have cleaning circuits for treatment of fines.

Jan 1, 1980

By R. J. Mechin

Top-slicing was introduced in the Pachuca district in 1917 by T. C. Baker, at that time mine superintendent of the Santa Gertmdis mine. There then existed 1200 ft. below the surface, lying between the fourteenth and eighteenth levels (a vertical distance of 354.2 ft.) a block of ore having a length of 295 ft. and an average width of 82 ft. and dipping at an angle of 65°. This block lay in the hanging wall of the vein and consisted of highly altered and fractured andesite, cut through with mineralized quartz stringers. On the foot-wall side, 26 ft. of old filling, left by former operators, was of payable grade. Stoping therein was first undertaken by the square-set filling system, but because of the extremely heavy hanging wall and loose dry filling in the foot wall, all square setting had to be accompanied by spiling, which made the operation exceedingly slow and costly. A system of mining was sought that would eliminate the difficulties inherent to the square-set system in heavy ground and that at the same time would effect the desired economies in mining. After several years, top-slicing has met all the conditions most satisfactorily, so that when similar conditions presented themselves in El Bordo mine, this method was adopted with necessary modifications. El Bordo mine is operated by the Santa Gertrudis company, and is situated about two miles north of Pachuca. Because of the diversity of the orebodies, a number of different mining methods are in use, among them top-slicing. Description of Block of Ore Former operators, when mining the original vein, left in the foot wall ore of too low grade to be profitable at the time; also, considerabe low-grade material was sorted out and stored as filling in the stopes. Some pillars were left in place, either for support or because they were not payable. All this has formed a body of ore that can now be exploited

Jan 1, 1925

By F. F. Hintze

Petroleum production in the Rocky Mountain district during 1929 registered a small decline from that of the previous year, conforming with a gradual decrease in the amount of oil produced during the last few years. A number of causes may be mentioned to account for this result, chiefly overproduction in other regions and the apparent desire of operators in this territory to conform to the policy of proration and retrenchment adopted by special agreement among operators in many of the fields. Unit operation has been adopted to control oil production from the Sundance sands in the Salt Creek field, Wyoming, in line with the general effort to stabilize this phase of the industry and prevent excessive drilling of new wells. Another factor which greatly retarded development in the Rocky Mountain region was the executive order of President Hoover issued in March refusing the award of new prospecting permits to future applicants and threatening cancellation of all existing permits upon which actual drilling had not been done as specified by law. Only four states in the Rocky Mountain region are important oil producers. In 1929 Wyoming led with an average daily output of 52,600 bbl., followed by Montana, with 10,730 bbl.; Colorado, with 6500 bbl., and New Mexico with 2700 bbl. daily average. These states bear the following percentage relation to total production in the Rocky Mountain district: Wyoming, 72; Montana, 15; Colorado, 9; New Mexico, 4., (Table 1.) The total average daily production was approximately 73,000 bbl. and the total production of all four states during the year was 26,534,606 bbl. This is 2.7 per cent. of the total oil production of the United States during this period. Colorado Although Colorado is not a large producer of oil and natural gas, its oil operations are widely distributed over the state. The Florence field is one of the oldest oil fields in the country, and there are good reasons for believing that the state will retain its place for many years to

Jan 1, 1930

By A. L. Sweetser

Experience has shown how difficult it is to obtain information regarding laboratory-tests in connection with the chlorina-tion-process for the extraction of gold from its ores, and I therefore present the following method, somewhat in detail, for the benefit of those who may desire to pursue research work in this field. The ore chosen was a partly-decomposed porphyry, extremely siliceous and comparatively soft, and of an average value per ton of from $25 to $26 in gold and $0.68 in silver. The chemical analysis gave: Fe,O,, 3.35; FeS, 1.5; MnO,, 0.75; ZuCO,, 4; A1,0,, 3.20; H,O, 0.7 ; SiO, and insol., 83.50 per cent. The ore, first crushed to pass a 5-mesh screen, was coned and quartered until a 45-lb. sample was obtained. Samples for assay and analysis, taken from this lot by means of riffles, were crushed to pass a 100-mesh screen. The original sample was divided into 10-lb. lots, of which one was crushed to pass a 10-mesh screen; another through a 12-; the third through a 20-, and the last through a 30-mesh screen. Because of the production of an excessive quantity of fine material it was deemed inexpedient to crush finer than 30-mesh size—a decision the wisdom of which is proved by the data in Table I., showing the increasing ratio of slimes, in proportion as the ore is crushed finer. In order to determine the percentage of fines in the sizes of mesh selected, a weighed sample of ore was taken from each mesh, and a sample from the 10-mesh lot was passed successively through 20-, 40- and 100-mesh screens, and the percentage remaining on each screen determined. This process was repeated with the 12-, 20- and 30-mesh lots, with the results given in Table I.

Jan 1, 1908

US 4,133,866-Selective recovery of the bound sodium content of red mud obtained in the production of alumina by the Bayer process The red mud is mixed with aqueous ferric sulfate, the resulting suspension stirred intensively at a temperature of 15-85°C and adjusted to a pH value of from 4.4 to 4 6, and the solids are removed. The clarified solution of sodium sulfate can be used directly for the leaching of additional bauxite T Lakatos, M Miskei, J Szolnoki, F. Toth, and L Revesz, assigned to Aluminiumipan Tervezo es Kutato Intezet, Almasfuzitoi Timfoldgyar, and MTA Geokemiai Kutatolaboratorium. Jan. 9, 1979. 3 pp (423-208). Filed: 10¬28-77 (846,588) Priority: Hungary, 10-29-76 (AU 364). US 4,145,398-Continuous digestion of bauxite with aqueous sodium hydroxide, with higher throughput in existing plants or the construction of smaller plants with equivalent capacity The ore suspension at a solids concentration of 1-6% by weight and temperature of 200-300°C is flowed through a digestion tube or nest of tubes at a velocity of preferably from 2 to 5 m/s and under a pressure of from 10 to 200 atm Heat is supplied by a fluid heat carrier flowing in countercurrent in a concentric outer tube. Preferably, some of the red mud is recycled to the digestion operation. L. Plass, assigned to Vereinigte Aluminium-Werke A.G Mar 20, 1979. 5 pp. (423-121) Filed: 2-27-76 (754,749) Same: British 1,437,548, dated May 26, 1976 Same Canadian 1,008,253, dated Apr 12, 1977. US 4,146,573-An aqueous slurry of red mud, formed in the Bayer process production of alumina from bauxite, can be solidified by treating the red mud with a described quaternary ammonium salt, continuing mixing until the red mud has set up to a rigid mass, and recovering the solidified concentrated red mud J Kane, assigned to Nalco Chemical Co Mar 27, 1979 4 pp (423-82). Filed 8-2-78 (930,251).

Jan 1, 1980

By W. Mostowitsh

I. Introductory LEAD sulphate occurs as anglesite, and is formed in every roasting of lead sulphides or sulpho-salts containing lead. In smelting in the blast furnace an ore containing natural or artificial lead sulphate, a large part of the sulphur present goes to the formation of the intermediary product matte which requires further treatment for the recovery of the values it contains. It is therefore of importance to study the behavior of lead sulphate at elevated temperatures, both without and with reducing agents, especially as the information available gives data which are more or less conflicting. II. Decomposition of Lead Sulphate by Heat The dissociation of PbSO4 into PbO and S03 or S02 and 0 has been discussed by many metallurgists. The following data are considered to be representative. Doeltz and Graumann1 found that PbSO4 was not affected by heat at 800° C.; that dissociation began at 900° C., and was rapid at 1,000° when the salt lost in 1 ½ hr. 14.3 per cent. of its weight, the dry salt containing 26.43 per cent. SO3. Hofman and Wanjukow2 summarize their work upon the dissociation of metallic sulphates with the statement that the decomposition of PbSO4 begins at 195° C., but progresses very slowly, with the formation of the basic salt 6PbO.5S03 at 705°; that this salt undergoes a transformation at 847°, begins to be decomposed at 888°, to sinter at 896°, and to fuse at 910°; rapid dissociation begins at 952°, and is accompanied by volatilization of PbO; retardations at 958° and 962° indicate an acceleration of decomposition which, however, is not complete. Proske3 found that PbSO4 gave off at 900° C. in 30 min. only 0.63 per cent. of its SO3, and that the loss increased with

Jan 5, 1916

By H. B. Menardi

The history, geology, ore deposits and current mining operations of the Huitzuco district have been described by C. W. Vaupelll and the current mill operations by David Segura.2 This paper completes the story of Huitzuco with a description of the reduction of the flotation concentrates. The predominant ore mineral in the Huitzuco deposit is livingstonite, a rare, mercury-antimony sulphide (HgS.2Sb2S3). Other associated minerals include cinnabar, stibnite and elemental sulphur. The gangue is largely limestone and gypsum. Former operators mined a selected grade of ore, from which the mercury was recovered by an expensive, inefficient operation. A considerable tonnage of rejects and fills, carrying mercury and antimony, was left in the mine. In order to recover both mercury and antimony from these fills and also from newly developed ore, a combination method has been developed, by means of which the livingstonite is concentrated by flotation and the flotation concentrates subsequently reduced to metallic mercury and antimony. An essential part of the flotation treatment, as developed by Segura, is the use of copper sulphate as an activating agent. A typical analysis of concentrates now being produced at the property is as follows: Hg, 7.73 per cent; Sb, 26.40; Fe, 1.20; S, 29.68; CaO, 11.50; MgO, 7.25; SiOz, 1.63; A1203, 0.65. PRocess foR Treatment op Concentrates In the development of a process for treatment of these concentrates, three different methods were investigated: (1) roasting in a Dwight-Lloyd sintering machine with an excess of air; (2) separation of mercury and antimony by distillation in a reducing atmosphere; (3) a third method, which ultimately developed into the present process. The third method involved a slow calcination at a closely regulated temperature with a limited admission of air, which separated the mercury and antimony, the raeroury passing out of the kiln with the furnace gases and the antimony remaining in the calcine or "sinter."

Jan 1, 1944

By Percival Roberts

I DESIRE to present this evening a few notes from the standpoint of a manufacturer upon the subject of tests of iron for structural purposes; not that I have any new facts to which to call your attention, but rather a statement of a few of the objectionable features of the present modes of operation, showing the absolute necessity of important and very radical changes. At the present time, as most of you are well aware, proposals for materials are accompanied by certain specifications as to quality and tests to which the manufactured article must conform. Evolved in many cases from the inner consciousness of some one more expert in theory than in practice, we find a few of the specifications to be good, many bad, and some indifferent. Appended are specimens of specifications of the most prominent engineers and railroad companies, which may be taken as representing the best and most intelligently framed specifications which we now have, arid yet in some pints scarcely satisfactory, owing to misunderstanding of tests as now made. To show the very varied opinions held as to proper quality of iron for bridge-work, I give the tests of two lots of fron made by inspectors with the results, which were referred to the maker with the following commensts : Lot No. 1. Teat of Iron for Eyebars. Reduced area. Elongation. Size. Per cent. Per cent. Breaking load. 4" x 1 3/8 26 25 68,500 Bent 120° around 1 1/2" bar with fracture. Reduced area. Elongation. Size. Percent. Percent. Breaking load. 6" x 1 7/15 24 24 48,820 Bent 110° around 2" bar. This iron should bend 180° without fracture, and is therefore rejected. Lot No. 2. Test of Iron for Brirlye. All measurements taken upon a parallel section of 10". Iron for tension members. Breaking load Elongation. size. per aquare inch. Percent. 1 5/8 49,600 20 14 61,600 27

Jan 1, 1882

By R. V. Whetsel, Valentin R. Garfias

Mexican petroleum production for 1928 was approximately 50,000,-000 bbl., the lowest in 12 years, showing a decrease from 1927 of over 14,000,000 bbl. The steady downward trend for the past 7 years has allowed Russia to outrank Mexico by a wide margin for third position and if the expected decline continues during 1929 it should place Persia ahead of Mexico as fourth in rank. The year 1928 was characterized by a continuation of the reentrench-ment previously begun by the oil industry in Mexico in an attempt to adjust itself to the adverse conditions caused by the steady depletion of the fields, the rapidly increasing competition from Venezuela and the United States and the as yet unsettled petroleum legislation and taxation situation. With the exception of the Isthmus of Tehuantepee, drilling in proved fields was considerably curtailed and although the number of wildcats completed exceeded those of the previous year, no extensive drilling programs were undertaken. Practically without exception, the major companies reduced personnel and operations to a point where they could meet existing conditions and mark time while awaiting a more stable basis for the future testing of unexploited regions. Many terminal and transportation facilities were abandoned during the year. Development Only 371 wells were completed during 1928, as against 583 in the previous year and 808 during 1926, and although the percentage of producers in 1928 was 34 per cent., compared to 26 per cent. in 1927, the initial production of -wells was considerably lower than during the previous year, a great number of the wells showing salt water and becoming "strippers" from the time they started to flow. Development is summarized in Table 1. Drilling in the Northern fields did not prove any new pools of importance, the most successful wells being located in the old producing areas of Cacalilao, Panuco and Topila. (Fig. 1 Very few of the producers made more than 1000 bbl. daily, and a great many less than 300 bbl. a day.

Jan 1, 1929

By Ellsworth Daggett

THE ore smelted in the Winnamuck furnace during the year 1872 consisted, for the most part, of oxidized ores from the Winnamuck mine, only sixty tons of outside ore (from the Spanish mine) having been smelted. The latter, like the principal Winnamuck ore, was oxidized or so-called carbonate ore. There was mixed with these oxidized ores three hundred to four hundred tons, or 7 to 10 per cent. of galena, some of which was mined with the oxidized ore, while a part was mined separately from the lower portion of the mine, and afterwards mixed with the ore, with a view of preventing the formation of deposits of metallic iron in the furnaces. The average assay in silver of all the ore handled was 51.46 oz. per ton, most of it existing as chloride of silver. The lead contents were 34.98 per cent., all, or nearly all, in the form of carbonate of lead. The relative amount of silver is not at all constant, the best silver ore often being poorest in lead. The predominant gangue was silica, several determinations of which have been made on representative samples, yielding in three such samples 26, 38, and 58 per cent. silica, the latter test being ore containing but little lead. The average contents in silica are about 35 per cent., with 6 to 7 per cent. sesquioxide of iron, and small quantities of alumina and lime. Mechanically, the ore was very fine, and so thoroughly disintegrated that it presented few distin¬guishing characteristics, rendering sorting or separating of ore from waste difficult, and often impracticable. Experiments on a small scale have been tried, with a view to separate the silica by washing; but these were unsuccessful, as the finest slime, requiring a long time to settle in still water, contained a large amount of silver; and on a careful sizing and washing of the sands and coarser parts, the silver contents were found to be less dependent upon specific gravity

Jan 1, 1874

By H. W. Seamon

The data here given were obtained by 1500 tests made by the United Verde Copper Co. to determine the most economical air pressure for the operation of hammer drills under the varying conditions of use, and to investigate the variation in drilling speed at different air pressures. Twelve models of drills were used at gage pressures ranging from 40 to 130 Ib. No effort was made to harmonize theory and practice; but rather to formulate empiric rules that would cover the average variation of the results obtained. These rules on the performance of hammer drills, based on the air pressure as the main variable, however, are not necessarily of universal application, but they apparently satisfy the results obtained in this series of tests. The drilling conditions at this property vary widely; an average of twenty-one machine shifts to a 3-ft. (0.9-m.) round has been found necessary in some of the development work, while an advance of 56 ft. has been made in seven shifts in the "oxide" ground. This range precludes the adoption of one type of drill as a standard; consequently almost every kind of hammer drill sold in this country has been tested. At present, sixteen models are in use, of which two types of the heavy (150 to 160 lb.) mounted drills, one of the light mounted drill, one stoper, and two hand plugging drills are considered as standard. Method of Conducting Tests All machines were of the wet type, and were selected as being representative of the several types in use at this property; all had been in use from two to four months except the Waugh 66 and the Sullivan DX61, which were new. All machines were thoroughly overhauled before they were sent to the testing ground. Tests were made on the following machines: Ingersoll-Rand, No. 248 Leyner-Ingersoll, No. 18 Leyner-Ingersoll, No. 26 Baby Leyner, BCRW430 Jackhamer, and CCW11

Jan 1, 1922

By Ellsworth Daggett

The ore smelted in the Winnamuck furnace during the year 1872 consisted, for the most part, of oxidized ores from the Winnamuck mine, only sixty tons of outside ore (from the Spanish mine) having been smelted. The latter, like the principal Winnamuck ore, was oxidized or so-called carbonate ore. There was mixed with these oxidized ores three hundred to four hundred tons, or 7 to 10 per cent. of galena, some of which was mined with the oxidized ore, while a part was mined separately from the lower portion of the mine, and afterwards mixed with the ore, with a view of preventing the formation of deposits of metallic iron in the furnaces. The average assay in silver of all the ore handled was 51.46 oz. per ton, most of it existing as chloride of silver. The lead contents were 34.98 per cent,, all, or nearly all, in the form of carbonate of lead. The relative amount of silver is not at all constant, the best silver ore often being poorest in lead. The predominant gangue was silica, several determinations of which have been made on representative samples, yielding in three such samples 26, 38, and 58 per cent. silica, the latter test being ore containing but little lead. The average contents in silica are about 35 per cent., with 6 to 7 per cent. sesquioxide of iron, and small quantities of alumina and lime. Mechanically, the ore was very fine, and so thoroughly disintegrated that it presented few distinguishing characteristics, rendering sorting or separating of ore from waste difficult, and often impracticable. Experiments on a small scale have been tried, with a view to separate the silica by washing; but these were unsuccessful, as the finest slime, requiring a long time to settle in still water, contained a large amount of silver; and on a careful sizing and washing of the sands and coarser parts, the silver contents were found to be less dependent upon specific gravity

By W. Mostowitsch

Lead sulphate occurs as anglesite, and is formed in every roasting of lead sulphides or sulpho-salts containing lead. In smelting in the blast furnace an ore containing natural or artificial lead sulphate, a large part of the sulphur present goes to the formation of the intermediary product matte which requires further treatment for the recovery of the values it contains. It is therefore of importance to study the behavior of lead sulphate at elevated temperatures, both without and with reducing agents, especially as the information available gives data which are more or less conflicting. 11. Decomposition of PbS04 by Heat The dissociation of PbSO4 into PbO and SO3 or SO2 and 0 has been discussed by many metallurgists. The following data are considered to be representative. Doeltz and Graumannl found that PbS04 was not affected-by heat at 800" C.; that dissociation began at 900" C., and was rapid at 1,000° when the salt lost in 1 1/2 hr. 14.3 per cent. of its weight, the dry salt containing 26.43 per cent. SO3. Hofman and Wanjukow2 summarize their work upon the dissociation of metallic sulphates with the statement that the decomposition of PbS04 begins at 195' C., but progresses very slowly, with the formation of the basic salt 6PbO.5SO3 at 705"; that this salt undergoes a transformation at 847") begins to be decomposed at 888") to sinter at 896", and to fuse at 910"; rapid dissociation begins at 952") and is accompanied by volatilization of PbO; retardations at 958" and 962" indicate an acceleration of decomposition which, however, is not complete. Proske3 found that PbSO4 gave off at 900" C. in 30 min. only 0.63 per cent. of its 803, and that the loss increased with

Jan 1, 1917

By T. Kunitake, H. W. Paxton

The self-diffusion coefficient of chromium in various alloys in the iron-chromium system has been measured. A variation in Dofrom 10-4 for pure chromium to a maximum of 102 near 60 pct Cr appears with a range of activation energy from about 50 to 80 kcal per mol. The general pattern of behavior observed is not consistent with present theories of diffusion by a vacancy mechanism. Suggestions are made for further clarifying experiments. RECENT work on the self-diffusion of chromium,' ?-uranium2-4 and ß-zirconium,23, 26 has indicated that these elements have very low Do (10-3 to 10-4) compared to that conventionally expected for vacancy diffusion. Pound et al.5 have suggested that this may be due to ring diffusion being operative, and have proposed a model which gives Do of the correct order of magnitude. The essential difference between this model and Zener's earlier model6 is that whereas Zener assumes tight coupling between atoms in a ring to establish a vibrational frequency, Pound's model allows the atoms to vibrate essen- tially each in their own normal modes. The probability of each atom moving in the correct direction to cause a ring of n atoms to rotate by 2p/n must then be calculated as a "synchronal entropy" which may be -20 e.u. for a 4 atom ring. In iron, the latest determinations of self-diffusion yield Do in the range 18 to 522.7-9 The most accurate is probably that of Borg and Birchenall8 which is 118 ± 3. This value is not inconsistent with that expected for a vacancy mechanism. Since iron and chromium are very closely similar in size, are both transition metals, and their solution is thermodynamically ideal at temperatures of interest in diffusion, it appeared that a determination of self-diffusion coefficients, Do and Q as a function of composition would be of substantial interest to diffusion theory. The following measurements were made with this purpose in mind.

Jan 1, 1961

By Chung Yu Wang

During the time I was in charge of this mine, from 1914 to 1915, I had occasion to read the interesting papers by T. T. Read and C. M. Weld about these deposits, to find how far their observations corroborated mine. To quote from the paper on "The Mineral Production and Resources of China" by T. T. Read:' "The iron ores at Tayeh * * * * lie along the contact between a marble and an intrusive body of a dark gray syenitic rock. * * * * At one place it is slightly magnetic, apparently having been partly reduced to the magnetic oxide by the action of reducing solutions, which have deposited small amounts of copper and iron sulphides along the foot wall." To quote from the paper on "The Tayeh Iron Ore Deposits"2 by C. M. Weld: "I find the orebodies described as occurring along an approximately E-W contact between hornblende-granite and limestone. * * * * The granite is a granitoid rock consisting essentially of quartz, potash-feldspar and hornblende."' These quotations show the differences of opinion entertained by two observers. My own study of these interesting deposits has been intermittent amid the manifold duties devolving upon a superintendent. General DescriptLon of the Deposits The deposits are situated to the west of the Yangtse River, being connected with the river by a standard-gage railroad of 66 li (18.16 miles, 26 km.) in length, as shown in Fig. 1. The general topography of the region is of mature age, showing broad valleys with sluggishly meandering streams. These deposits are worked as two mines, named the Tieh Shan Mining Department and the Teh Tao Wan Mining Department; the former consisting of three orebodies more or less connected, named Tieh Man Kan, Sha Mau Tze and Lung Tung, and the latter consisting of Chang Pei Shan, Sze Tze Shan, Ta Shih Men, Yeh Chih Ping and Tsin Shan, also more or less connected, as shown in Fig. 2.

Jan 1, 1918

By Chung Yu Wang

DURING the time I was in charge, of this mine, from 1914 to 1915, I had occasion to read the interesting papers by T. T. Read and C. M. Weld about these deposits, to find how far their observations corroborated mine. To quote from the paper on "The Mineral Production and Resources of China" by T. T. Read:' "The iron ores at Tayeh lie along the contact between a marble and an intrusive body of a dark gray syenitic rock. * * * * At one place it is slightly magnetic, apparently having been partly reduced to the magnetic oxide by the action of reducing solutions, which have deposited small amounts of copper and iron sulphides along the foot wall." To quote from the paper on "The Tayeh Iron Ore Deposits"2 by C. M. Weld: "I find the orebodies described as occurring along an approximately E-W contact between horn-blende-granite and limestone. * * * * The granite is a granitoid rock consisting essentially of quartz, potash-feldspar and hornblende." These quotations show the differences of opinion entertained by two observers. My own study of these interesting deposits has been intermittent amid the manifold duties devolving upon a superintendent. General Description of the Deposits The deposits are situated to the west of the Yangtse River, being con¬nected with the river by a standard-gage railroad of 56 li (16.16 miles, 26 km.) in length, as shown in Fig. 1. The general topography of the region is of mature age, showing broad valleys with sluggishly meandering streams. These deposits are worked as two mines, named the Tieh Shan Mining Department and the Teh Tao Wan Mining Department; the former consisting of three orebodies more or less connected, named Tieh Man Kan, Sha Mau Tze and Lung Tung, and the latter consisting of Chang Pei Shan, Sze Tze Shan, Ta Shih Men, Yeh Chih Ping and Tsin Shan, also more or less connected, as shown in Fig. 2.

Jan 3, 1917

By Wheeler O. North

The United Eastern Mining Co.'s property is in the Oatman, Gold Roads mining district of Mohave County, Ariz. The mine and mill are 26 mi. (41.8 km.) southwest of Kingman, the nearest railway connection. The ore consists of a mixture of calcite and quartz with some andesite. The values, which are mainly gold, alloyed with approximately 34 per cent., by weight, of silver, are extremely fine so that fine grinding is a most important factor in extraction. The milling process, briefly, consists of single-stage, coarse-crushing, two-stage ball milling in cyanide solution, followed by combined mechanical and air agitation, and by straight counter-current decantation. The Merrill zinc-dust precipitation method is used to recover the values, and is preceded by the Crowe vacuum treatment of the solution. The most notable features of the plant are the absence of filters, and the use of short peb mills loaded with steel balls, for fine grinding. As originally designed, the capacity of the mill was to be 200 tons per day, although the grinding end was to be capable of handling double that amount. By adding extra agitators and thickeners and by some alterations to the grinding end, the daily capacity has been raised to 285 tons. This, it is hoped, may be increased to 300 tons during the warm months of the present year. Tonnage Milled per Day 1917 average....... 234; maximum month.... 272; minimum month..... 146 1918 average....... 253; maximum month,... 276; minimum month..... 217 As this article primarily deals with plant operations, those interested in the mechanical description are referred to the article by Mr. Otto Wartenweiler,1 the designing engineer, which gives the details and costs of design and construction. The following changes have been made since that article was published: Two special simplex Dorr classifiers have been substituted for the two duplex Callow screens and bucket elevator, which classified the Marcy discharge and returned the oversize for regrinding. Three Dorr

Jan 1, 1920

By Howard N. Eavenbon

The low-volatile, or smokeless, coal field of southern West Virginia is in Fayette, Raleigh, Wyoming, Mercer, Summers and McDowell counties, in the extreme southern portion of the state, and extends into the adjacent counties of Tazewell and Buchanan in Virginia. McDowell County is the largest producer, with Fayette and Raleigh counties next. The total area of the field (Fig. 1) is approximately 1600 square miles, or 1,024,000 acres. The percentage of volatile matter in the various seams ranges from 15 to 26 per cent, The upper limit of true low-volatile coal is usually considered to be 23 per cent volatile matter; above this point, the coal usually is classed as medium volatile. This limit will be used in this paper. Geological Position of Seams The smokeless seams are among the oldest coal seams in the country, as the Pocahontas group of rocks is the lowest member of the coal-bearing series and rests on top of the Mauch Chunk red shale, in or below which no coal seams of any importance occur. The Pocahontas group, or Lower Pottsville measures, has a thickness of 750 ft. in southern McDowell County, which decreases to about 350 ft. at Prince, on the New River, in Fayette County, 50 miles to the northeast. In this group, at various points, 12 coal beds have been found, of which 4 have been found in workable thickness and are now being mined. The New River group, or Middle Pottsville measures, is just above the Pocahontas group and varies in thickness from 1300 ft. in southern McDowell County to about 700 ft. along New River. At various places, 16 beds of smokeless coal have been identified in this group, of which 5 have been found in workable thickness and are now being mined. A general columnar section of these strata is shown on Fig. 2. The workable seams in these strata, arranged in geological sequence, are as shown on page 77.

Jan 1, 1932

By Samuel B. Christy

Columbia College has the honor of founding the first well-organized School of Mines in America. The University of Michigan, however, shared with her the honor of graduating the first class, in 1867. Prof. Thomas Egleston was the pioneer in the new movement. He drew up the original plan for the Columbia School of Mines in 1863, and was in 1864 appointed its first professor. The Massachusetts Institute of Technology turned out a class of mining students a year later, and others followed in the order of their first graduating class, as shown in Table I., condensed from the Engineering News of August 11, 1892. Besides those mentioned, mining schools have been organized in the States of Arizona and Nevada, and more or less instruction is advertised in connection with other engineering schools. It will be noticed that Columbia easily heads the list with reference to the number of graduates; the Massachusetts Institute of Technology stands second ; the California University third, and then follow Lehigh, Washington University, the University of Michigan, Lafayette College, Michigan Mining School, Missouri, Colorado and the others. The total number of graduates from these 16 schools in 26 years, has been 871, or an average of 33 per year; but as all these schools give a partial education to many who do not graduate, it is probable that from 2000 to 3000 mining students have received from them considerable training. An interesting comparison of the rate of growth of the engineering schools has been recently instituted by the Engineering News, for the period between 1860 and 1892. Table 11. shows that civil engineering had the advantage in numbers from the start, and maintained the advantage until 2890, when it was surpassed by mechanical engineering. The development of the latter course may be said to date from 1880. There are many interesting facts evident in Table 11. It will

Jan 1, 1894