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By G. T. Brookman, P. B. Katz, D. K. Martin, J. A. Ripp

With the increasing emphasis on using American fossil fuel resources, more facilities are evaluating coal conversion. In addition, those facilities now using coal are considering better coal storage procedures to minimize negative environmental impacts. For example, rainfall and snowmelt on a coal pile produce a drainage stream high in acidity, suspended and dissolved solids, iron, sulfate, and trace metals. Usually, the characteristics of the runoff stream do not allow for direct discharge to streams and lakes. Detention in a holding pond and pH adjustment are traditional methods of treating coal pile runoff before discharge. Historically, this treatment methodology has been performed inefficiently without a good data base to assess how much runoff treatment is needed to meet effluent guidelines. A model was developed to simulate runoff volume and pollutant loadings under different rainfall conditions. This model will be used to size and design treatment works for coal pile drainage. The Ohio State University (OSU) version of the Stanford watershed model was selected as the base model. The OSU model simulates overland runoff in a rural area. It includes factors such as soil moisture, depression storage, infiltration, and flow to a ground water reservoir. TRC also used the OSU co-model for acid production in coal refuse piles as the basis for qualitative modeling work. A number of modifications were made to the OSU models to make them applicable to coal storage piles. The physical/chemical reactions included in the models are described in more detail in the project report. In addition, the models were altered so that a standard NOAA magnetic tape of historical meteorological data can be used as data input instead of requiring the user to create a data file. The resulting TRC coal pile drainage model is actually two models - a quantitative model, termed H2OTRC, and a qualitative model, TRCCOAL.

Jan 1, 1984

By Eugene A. Stephenson

THE year 1935 is noteworthy for the superior legislative measures passed by several states. Sincere efforts have been made to establish equity between the various producing tracts in oil and gas fields. Outstanding legislative measures include the New Mexico law, known as Senate Bill No. 79, approved on Feb. 23; the Oklahoma law on April 22; the Texas law on April 30; and the Kansas law, which became effective July 1. Primarily designed to prevent waste, but more or less the direct out- growth of the current excess of potential oil and gas production, these statutes fall in the class of "Proration Laws."

Jan 1, 1935

By R. V. Whetsel, V. R. Garfias

The production of petroleum in Mexico during 1933 totaled 33,430,-000 bbl., or 625,000 bbl. more than in the previous year. This is the first time since the peak in 1921 that the declining trend of output has been reversed. Drilling activity in the proven areas of the Isthmus of Tehuantepec and the Northern fields increased substantially during the year. The completion of two producers in the Poza Rica area south of Tuxpam and a shallow gas well in the Camargo district near the American border were the outstanding features of development. Late in September the Northern fields and the oil port of Tampico were swept by a hurricane followed by floods which caused loss of life and heavy damage to the City of Tampico and to transportation, tank farms and refineries. For three weeks, owing to the silting of the Tampico Bar and destruction of terminal facilities, no oil could be shipped for export. About half the wells of the Northern fields were closed for a period of 30 days. The prompt and effective steps taken by General Macias, chief of military operations, prevented the spread of epidemics and any serious aftermath of the catastrophe.

Jan 1, 1934

By L. R. Plitt, M. K. Kim

The collecting properties of the fatty acid type collectors for barite were studied using zeta potential measurements, infrared spectroscopy, and Hallimond tube flotation tests. Based upon the experimental results, an adsorption model has been formulated which emphasizes the predominant role of the hydrocarbon chain in collector adsorption.

Jan 1, 1975

By Werner Goldsmith, Akihiko Kumano

An experimental investigation was conducted to study the response of soft, porous rock, green shale, under dry and liquid-filled conditions to normal impact of hemispherically-tipped cylindrical steel projectiles with diameter and mass of 6.35 mm and 31 g, respectively. Disk-shaped green shale specimens with nominal thickness and diameter of 32 mm and 127 mm, respectively, were tested in the initial projectile velocity range of 18 to 42 m/s. For the liquid-filled tests, the green shale specimens were saturated with kerosene. A total of eight impact tests were conducted. Information on the extent and pattern of the damage in the rock and velocity histories of the projectiles were collected. It was found that while the penetration depth was quadratically related to the initial projectile kinetic energy for the dry case, the penetration depth for the liquid-filled case was directly proportional to the initial projectile kinetic energy. The results of the impact tests under dry and liquid-filled conditions conformed well to a linearly viscous model and a Coulomb frictional model, respectively. The major deformation processes in the dry specimens were found to be compaction and equivolumnial distortion of the material. In the liquid-filled case, however, formation of cracks and equivolumnial distortion with less material compaction were observed. Furthermore, craters were formed as a result of the crack formation in the liquid-filled specimens.

Jan 1, 1982

By William Ian Wark, Alwyn Birchmore Cox

DIFFERENT parts of the world in which mineral separation by flotation is practiced experience vastly different natural temperatures, and in some districts there is a big difference between summer and winter temperatures. It will be shown in this paper that the temperature is of considerable importance where it is desired to separate pyrite or arsenopyrite from other minerals. It is probable that some of the differences obtained in the flotation of these minerals at plants working under otherwise similar conditions are due to temperature variations. The influence of temperature on the flotation of galena, chalcopyrite and sphalerite by xanthates is less marked than its influence on pyrite flotation. EXPERIMENTAL METHOD The methods used have been described previously.1-4 In each test the clean mineral specimen was immersed in a solution prepared by adding the reagents in the following order: (1) sodium carbonate and hydrochloric acid as necessary to buffer the solution and to give approximately the pH value desired, (2) sodium cyanide, (3) copper sulphate, (4) ethyl xanthate, followed by (5) a final adjustment of pH value. After ½ hr. the mineral was tested for a xanthate film by bringing a bubble of air into apparent contact with the submerged polished surface. If true contact (spreading) was not obtained within ½ min. and after cleaning the surface by wiping with a small linen pad, it was concluded that the adsorption of a xanthate film had been prevented. Following the method already described (ref. 2), curves have been constructed such that for a solution ' composition corresponding to any point below the curve, contact between mineral and air is possible, whereas for a solution composition corresponding to any point above the curve, contact is impossible.

Jan 1, 1938

Jan 1, 1940

"The Washoe Smelter is situated about two miles east of the City of Anaconda. The smelter site includes about 240 acres and peculiarly adapts itself in topography to the efficient handling of material. By reason of its being on a hill side it is possible to make the level of delivery of the product of one building the level of the charge floor of the next succeeding one.THE ORES: The ore treated comes principally from the Butte district and is hauled by the Butte, Anaconda & Pacific railway, a distance of twenty-eight miles, in trains of fifty 50-ton bottom-dump cars to the railroad yards at East Anaconda. At this point these trains are• split up according to their destination,—that is, as to whether they go to the concentrator and sampling mill ore bins, or to the storage bins. In each case they are made up into trains of eight or sixteen cars, according to the size of engine to be used, and are carefully weighed before delivery to their respective bills.Some idea of the magnitude of the works may be obtained from the following figures:Amount of ore that can be treated in 24 hours…12,000 tonsLime rock from adjacent quarries…2,300 tonsCoke used…500 tonsCoal for reverberatory use…500 tonsCoal for power use…50 tonsWater, per minute, 35,000 gallons.Men employed in and around Anaconda, 3,000. Monthly payroll for above, $300,000.00."

Jan 1, 1913

Robert Annan is an Englishman who started his brilliant mining career in Arizona, yet achieved the pinnacle of success in his native land when he became chairman of The Consolidated Gold Fields of South Africa. This story of his life, prepared in collaboration with the late Henry Carlisle, continues this fascinating biographical series of important personalities in the mining industry.

Jan 12, 1964

By F. W. McQuiston

INTRODUCTION IDARADO Mining Company's mill and surface plant are at the portal of the Treasury tunnel at elevation 10,625 ft, 12 miles south of Ouray, Colo. In 1943 and 1944 this tunnel was extended in a westerly direction to tap the Black Bear vein, and, together with the recently completed 1100-ft inclined Treasury raise, provides access to the Black Bear mine. In 1945 a 250-ton mill was placed in operation to process development ores containing much needed copper, lead and zinc for war requirements, and to serve as a pilot testing unit for larger scale milling operations. In the years 1902 to 1927 some 329,000 tons of ore were produced from mine workings on the Black Bear vein that extended downward from elevation 12,325 ft at the Black Bear tunnel portal to elevation 11,618 ft, the lowest level of the mine. The ores were taken to the Cascade and Smuggler-Union mills near Telluride over two aerial tramways. Gold bullion, a copper-lead concentrate, and a zinc concentrate were produced. ORES The Black Bear is a compound fissure vein that has undergone several stages of mineralization. The first important stage is represented by complex ores of copper, lead and zinc in a quartz gangue. Precious metal content in this stage is low. The galena is argentiferous and varying amounts of tetrahedrite, freibergite and bornite tend to increase the normal copper content represented in chalcopyrite. The zinc occurs as a true sphalerite with little or no iron. Later re-openings of the fissure took place more or less co-extensively with the successive introduction of (1) abundant quartz gangue containing minor amounts of chalcopyrite, galena and sphalerite; (2) free gold in a fluoritic quartz gangue; and (3) barren, low-grade quartz with minor amounts of pyrite as the only important sulphide. The sulphides show practically no oxidation and little mineralogic intergrowth. Because of the common overlapping of these successive stages of mineralization within a given mining area, considerable dilution of the base-metal content takes place, thus lowering the average mill-head grade. PLANT PRACTICE Metallurgical testing was started shortly after the Black Bear vein was encountered on Treasury tunnel level in 1944. Laboratory tests showed good flotability of the minerals in the ore and indicated an economic separation could be made of the copper, lead and zinc minerals into their respective concentrates. The success of the separation is close reagent control to maintain a balance between promotion and depression of the sulphides. To maintain this balance reagent consumption for the three metal separation is considerably higher than in ordinary lead-zinc flotation.

Jan 1, 1947

By W. J. Whiten

Current developments in one type of crushing and screening model is reviewed and general design techniques and policies are proposed for the control of crushing plants.

Jan 1, 1984

By H. F. Mills

THE Iron King mine is 12 miles east of Prescott Arizona, in low foothills of the Agua Fria mining district. It was operated in 1906 and 1907, using gravity methods of concentration, mostly on oxide ores. In 1923 and 1924 shipments of sulphide ores were made to the Humboldt smelter. It remained inactive and stripped until 1936, when it was unwatered and sampled to the 400-ft level, the bottom at that time. In 1937 the Iron King Mining Co. purchased the mine and shipped oxide ores. During 1938 a bulk flotation plant of 140 tons daily capacity was erected. The mill has been enlarged and operations have been continuous since the mill first started. In July 1942, the Shattuck Denn Mining Corporation purchased the physical assets from the Iron King Mining Co., and since that time it has operated as the Iron King Branch of the Shattuck Denn Mining Corporation. Economic metals recovered are gold, zinc, silver, lead, a little copper, and fluxing pyrite. The ore occurs as a series of northward-dipping shoots arranged in echelon pattern, the series extending over a length of 3500 ft. Strike lengths of shoots vary from 60 to 600 ft, and widths from 3 ft to as much as 14 ft. The ore occurs in a sheared zone in pre-Cambrian chloritic schist. This schist belt is about 6 miles wide, and the ore zone is about 4 miles from the pre-Cambrian granites to the west. No intrusive rocks have been found in the mine, nor is there evidence of any on surface that might indicate a genetic connection with the ore occurrence (Fig I). The schists strike N23°E and dip 75° to the northwest, which might indicate an overturned thrusting from the granites. In the upper levels, the echelon pattern in which the ore occurs was quite regular, no disturbance of the enclosing schist, or of any faulting that might contribute to the formation of the ore pattern was noted. There is a possibility that the ore zone may be on the axis of an isoclinal fold. Another interesting feature is the more or less typical zoning of minerals along the strike in each of the ore shoots (Fig 2). The north end of each shoot is enclosed by schist, while the south ends usually extend out at full width beyond the economic mining limits in a pyrite-quartz low-grade material. The plane of demarcation between ore and schist is very sharp, with little or no impregnation of sulphide in the schist, especially on the footwall side. The footwall schists are usually firm and strong, and vary from dioritic phases to fine-grained, lighter colored bands, almost silky in texture. The hanging wall for a width of 50 to 8o ft consists of alternating sheetings of chloritic and pyritic schists. The mine is now partly developed to the 92o-ft level, with some diamond drilling from the 1150-ft station of No. 6 shaft. Economically there has been little change in nature or grade of ore from the top of the sulphides, about 200 ft below surface, to the bottom levels. There have been a few echelons in depth, just as there are in plan, and while a few shoots have had

Jan 1, 1947

By Peter H. Fletcher

THE ECONOMIC THEORY It will assist in the understanding of the Resource Rent Tax ("RRT") proposals in Australia if the economic theory behind the tax is briefly explained. The idea goes back to the concept of rent put forward by the nineteenth century economist David Ricardo. His analysis was broadly as follows: A tenant who farms land which is just on the margin of economic viability cannot afford to pay his landlord any rent, because the value of the crop produced by such marginal land is equal to the cost of producing the crop, plus the cost of keeping the tenant alive. There is nothing left out of which to pay rent to landlord. A tenant who farms good land can afford to pay his landlord rent -- but the maximum amount of rent that he can be asked to pay is the excess of the value of the crop produced by the good land over the cost of producing the crop and of keeping the tenant alive. Adapting this theory to the taxation of mining projects, modern economists who favour this system of taxation say that a profitable mining (or resource) project produces rent in the Ricardian sense. The amount of the rent is the excess of the profits by the project over the amount of profit which is required to attract the taxpayer to invest in the project in the first place. This explains why the tax is called "Resource Rent Tax". PAPUA NEW GUINEA PRECEDENT Although Papua New Guinea ("PNG") is now independent of Australia, there are still close links between the two countries, and strong similarities between their respective tax regimes. PNG introduced an Additional Profits Tax ("APT") in 1978. In many respects, Australia's RRT proposals follow the precedent set by PNG1s APT. However, in practical terms, not a great deal can be learned at this stage from PNG's experiences since, according to the writer's understanding, no PNG mining project has yet paid any APT. The Bougainville mine is not liable to APT, being subject to tax under its own separate Act, which contains a different formula. Outline of PNG Additional Profits Tax. PNG has two versions of Additional Profits Tax ("APT) - one applying to general mining, the other to petroleum mining. The two versions are very similar. The general mining version is described below. The tax applies on a licence by licence basis. This means that losses on Licence A cannot be offset against profits on Licence B. Expenditure relating to two or more licences, is apportioned "in such manner as the Chief Controller thinks reasonable". Tax is levied on the "accumulated value of net cash receipts" at the rate of (70 - n)% where n=the rate of income tax. The accumulated value of net cash receipts ("av") is calculated as follows: Calculations commence from the first year in which substantial expenditure was incurred on the construction of facilities for the carrying out of mining operations on the licence. For that year the av consists of the total allowable expenditure to date on that lease. For the second year, the av is the av at the end of year one indexed in accordance with the statutory formula set out below, plus the total allowable expenditure in year 2. For year 3, the av is the av at the end of year 2 adjusted by the index factors, plus the total allowable expenditure in year 3. This process is repeated year after year. When the licence starts to earn revenue, the revenue is deducted from total allowable expenditure for that year. If revenue exceeds total allowable

Jan 1, 1985

By Stewart K. Griffiths

A general method-of-lines numerical approach for modeling gas driven fractures has been developed and tested. The numerical solutions agree very well with known similarity solutions for laminar and turbulent flow. Acceptable accuracy is obtained with a very few numerical grid points and with modest execution times. The model has also been checked against the results of several field tests, in which air was used to drive axisymmetric fractures into permeable tunnel-bed tuff.

Jan 1, 1984

By E. C. IBBOTSON

THIS process was reported upon by the Canadian Commission in 1904, and much detailed information was also given in a paper by Chief Engineer V. Engelhardt.1 Believing that some of the latest particulars of the process will be of interest, I submit the following data. At Gysinge in Sweden during the year ending May 31, 1906, from a fixed furnace giving 1 ton (2,240 lb.) of steel per tap, there were produced 950 tons of tool-steel and special steel ingots. . In carbon and iron tool-steels all the usual tempers were made. The bulk of this steel was made from charges composed of about 80 per cent of Swedish white pig-iron and 20 per cent. of steel-scrap. The percentage of carbon was regulated by the addition of briquettes. Other charges were made from Swedish white iron and steel-scrap. The average time taken per charge, for the year, when adding briquettes, was 7* hr., and the electric energy consumed was 1,128 units (kilowatt-hours) per ton. The average time per charge for white-iron and scrap charges was 5.5 hr., and the electric energy consumed was 886 units per ton. These consumptions include all time and energy lost from various causes, such as bad water-supply, ice, etc. It is quite ordinary practice to complete the charges with briquettes in 6.5 hr., and without briquettes in 5 hr., as shown by the following typical charges White pig-iron, 1,457; steel-scrap, 439; briquettes, 220; ferrosilicon (50 per cent. of silicon), 17 ; ferromanganese (80 per cent. of manganese), 15 lb. ; and aluminum, 1 oz.=

Nov 1, 1906

By M. S. Tsai, Y. D. Chuang, H. M. Wu

To contact the commercial OPAP-kerosene solution with CH3OH – H20 solution, we used the appropriate concentration of DOPAP / MOPAP solvent. By using this treated OPAP solvent, uranium can be recovered from wet-process phosphoric acid. Electric DC power is supplied on the dual side of the ion-exchange membrane of a specially designed box-type mixer settler. U(VI) is reduced to U(IV) in the cathode and U(IV) is oxidized to U(VI) in the anode. Uranium is extracted in the cathode by treating it with OPAP solvent. Next the uranium is stripped from the loaded OPAP solvent by concentrated phosphoric acid in the anode. After 300 hours of continuous operation total recovery of uranium is as much as 95%.

Jan 1, 1984

By Jacob H. Masliyah

The dynamic simulation of a gravity separation vessel (a high throughput classifier) has been developed. The feed slurry was assumed to contain light (oil droplets) and heavy (mineral solids) particulates of various size. The feed slurry was characterized by the particle size distribution of both the light and the heavy particles. The equations describing the slip velocity for each specie and those defining the time variation of each specie concentration were solved implicitly using Runge-Kutta algorithm. The dynamic simulation was capable of predicting the movement of the froth layer and the density variation of the under- flow. Field data are in agreement with the simulation model and it is shown how this model could be used to develop a control strategy.

Jan 1, 1984

By Francis Laur

BAUXITE, at first considered as a mineralogical curiosity without importance, now attracts daily increased attention from mineralogists, geologists, and manufacturers. The metallurgy of aluminum, which has entered with the advent of electrolysis upon a new phase within the last few years, requires every day larger quantities of pure alumina, and this can only be obtained from bauxite, which is, of known substances, the richest in free alumina. Bauxite has been mined since 1872 in France, whereabout 200,000 tons have been produced, and where, as a natural consequence, its deposits, qualities, and applications have become known. Within a few years past, discoveries of important deposits of bauxite have been made in the United States and Canada, and this almost entirely new subject has thus acquired an additional interest for the mining engineers and metallurgists of the American continent. A brief summary of the available data concerning it appears likely, therefore, to be appropriate and valuable.* I.—HISTORICAL. It was in 1821 that the famous chemist Berthier discovered at Baux (Bouches du Rhône) a hydrate of alumina (varying from 66 to 79 per cent. in its contents of the oxide) mixed with silica and

Jan 1, 1895

By Y. L. Yao

It is suggested that the binary alloy systems of transition metals may be classified into groups according to "excess energy"E. For systems with positive ?E, the Laves phase, the TiNi3-, the TiCu3- and the Cudu-type rich in the smaller atom and the TizNi-type rich in the larger atom may exist. For systems with negative ?E and composed of two components of different crystal structures, the X, the Cr3O and the phase may exist on the side of the phase digrattz rich in the component with the higher "solubility factor" value am! the Cudu-type and the disordered hcp phase on the other side; if the crystal structure of two components is the same, the Cudu-type may exist on both sides. For systems with weakly positive ?E and cotnposed of different crystal structures, the u phase and the MgCd3-type may replace the Laves phase and the other types; the X, the Cr3O and the a phase may also exist. Finally, the equirnolar compounds of the CuAu- and the CsC1-type may exist irrespective of the sign of ?E. THERE are many factors which play a major role in the formation of binary intermediate phases in alloys, such as the crystal structure, the electronic structure, the atomic size and the electrochemical factor. To find the simultaneous effects of all these factors on the formation of intermediate phases of every metal would be a difficult task, especially because only a small fraction of experimental facts necessary for the complete formulation is known at present. Thus the past surveys of intermediate phases have been usually confined to few metals, several types of intermediate phases and some but not all controlling factors. In a previous paper1 the idea of "solubility factor" of metals was introduced. Although the suggested values are tentative and empirical, the formation of a continuous series of solid solutions in metals could be correlated with the interplay among the atomic size factor, the solubility factor and the crystal structure factor. The purpose of this paper is to demonstrate qualitatively the interplay of these factors in the formation of binary intermediate phases among transition metals. FACTORS CONSIDERED The Crystal Structure Factor. With rare earths excluded, the transition metals are classified into three common crystal structures listed in Table I. There is arbitrariness in the present classification of several allotropic elements. However, out of seventy-odd continuous solid solutions formed among transition metals, only six of them, namely, those

Jan 1, 1962

By J. W. Gilbert

DuriNG the sinking of the Needmore shaft, on one of the leases of the Eagle-Picher Mining and Smelting Co north of Webb City, Missouri, a very strong flow of water was encountered at a depth of 140 ft. Because OF the speed at which the water rose in the shaft after pumping was discontinued, and in order to speed up sinking operations, a new type of shaft bar and drilling equiprnent was devised, which saved approximately one hour on each shift This bar is made of an ordinary scren type shaft bar with a common drill arm attached to the bar by clamps and welded Into place To this drill arm is attached an ordinary arm clamp with set collar below Through this clamp is attached another drill bar of ordinary length so constructed that it slides back and forth and rotates around the perimeter of the shaft on the set collar A tripod saddle with legs removed is welded to the rotating bar, which enables the drill operator to drill holes at any angle completely around the shaft and also, after an arm extension, permits the drilling of the corner holes without removal of the rig from the saddle Upon completion of the drilling of a shaft round, all that is necessary for the complete removal of the drill machine and shaft bar is to engage the hoist hook through a. ¾-in. eye provided for that purpose on top of the bar and disconnect the air and water lines from the machine, thereby making it possible to pull the whole apparatus out in one piece. At no time is it necessary to remove the machine from the saddle Screw jacks are installed on the vertical bar to further ensure the rigidity of the shaft and bar arm. The accompanying sketch illustrates the details and method of setting up.

Jan 1, 1946