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By Philip Wilson

THE Province of British Columbia covers 382,000 sq. mi., about 250,000 sq. mi. of which have not been prospected. In fact, the coast country and the islands are so heavily timbered and the surface covered with so thick a mantle of soil, muskeg, and, over broad areas, glacial drift, that prospecting, and even trail breaking are extremely difficult. It is only along the coast or stream channels, or above the timber line, or where snow or rock slides have scoured the surface that rock outcrops can be found. Glaciers nestle in many of the cirques and fill the valleys in the higher ranges. Great ice sheets and perpetual snow persist in certain localities, thus reducing the exposed areas worth prospecting. Almost the entire country has been subjected to intense glaciation, so that the conspicuous oxidized outcrops of orebodies typical of the south-western United States and Mexico do not exist and surface bodies of enriched or oxidized ore are rare. With few exceptions, the mines1 of the Province are working primary ore. Developments have proved that some of these orebodies are large and important, although usually comparatively low in grade. Besides the few mines that have produced most of the minerals, there are many small mines, but their total output is unimportant. Considering the number of mines and prospects and the widespread mineralization, it is strange that no really great mines, comparable in size and richness with those in the western United States have been found.

Jan 8, 1922

By P Shenon

THE Elk City district is in north-central Idaho about 60 miles east of Grangeville and near the headwaters of the South Fork of the Clearwater River (Fig. 1). At the height of its boom in the early sixties, Elk City had a population of more than 2000 people.1 Published estimates of the placer production from the district range from $10,000,000 to $18,500,000. Although the first quartz-vein location was made as early as 1870, very little gold was extracted from quartz ores until after the completion of the American Eagle mill in 1902. Between 1902 and 1932 a total of from $750,000 to $1,000,000 was taken from about six properties.

Jan 1, 1934

By H. R. Tunnell

Every one connected with a mine knows that it is hard to keep down the costs of moving ore from the place where it is broken to the shaft or portal. Considered broadly, the subject of handling would cover all work done in a mine, but here it has been limited to handling ore in stopes and drifts, through chutes, and on levels. Good ventilation and a strict enforcement of rules for the prevention of dust safeguard the health of the men and lower transportation costs. Handling ORe in Stopes and Drifts The object of all stoping is to mine ore of good grade at low cost. In Butte, most of the ore is mined, from veins dipping from 60° to the horizontal to those which are vertical, by means of square-set, rill, timbered-rill, and back-filling stopes. When blasted, a part of the rock is carried by gravity to the grizzlies over the chutes, the rest is shoveled. Rock, when broken in a stope, may be deflected by inclined floors or slides toward the top of a chute leading to the level below. Ore broken clean enough to ship should be carried, by the force of its fall, into a chute or.as nearly into that chute as the safety of the stope and the men permits. Ore that requires sorting, about 80 per cent. of all ore mined, is caught on a grizzly and the waste picked out. Waste rock broken in a stope is used to fill that stope and should be blasted into the gob with as little handling as possible. About two-thirds of all the ore mined in Butte comes from square-set stopes; this method was used in nearly all cases until 1912, when the rill method of stoping was introduced in the Tramway mine, of the Anaconda Copper Mining Co. The shoveling and timbering costs in square-set stopes are high when compared with the costs in the rill stopes. All the ore cannot be carried to the chute, in falling from place, unless the chutes are very close together or the distance from the back of the stope to the top of the chute is sufficient to slide all the ore broken. The number of chutes is limited by the tendency of the ground to cave; if carried too close together the stope is weakened. Small stopes, or blocks,

Jan 1, 1923

By H. R. Tunnell

Every one connected with a mine knows that it is hard to keep down the costs of moving ore from the place where it is broken to the shaft or portal. Considered broadly, the subject of handling would cover all work done in a mine, but here it has been limited to handling ore in stopes and drifts, through chutes, and on levels. Good ventilation and a strict enforcement of rules for the prevention of dust safeguard the health of the men and lower transportation costs. Handling ORe in Stopes and Drifts The object of all stoping is to mine ore of good grade at low cost. In Butte, most of the ore is mined, from veins dipping from 60° to the horizontal to those which are vertical, by means of square-set, rill, timbered-rill, and back-filling stopes. When blasted, a part of the rock is carried by gravity to the grizzlies over the chutes, the rest is shoveled. Rock, when broken in a stope, may be deflected by inclined floors or slides toward the top of a chute leading to the level below. Ore broken clean enough to ship should be carried, by the force of its fall, into a chute or.as nearly into that chute as the safety of the stope and the men permits. Ore that requires sorting, about 80 per cent. of all ore mined, is caught on a grizzly and the waste picked out. Waste rock broken in a stope is used to fill that stope and should be blasted into the gob with as little handling as possible. About two-thirds of all the ore mined in Butte comes from square-set stopes; this method was used in nearly all cases until 1912, when the rill method of stoping was introduced in the Tramway mine, of the Anaconda Copper Mining Co. The shoveling and timbering costs in square-set stopes are high when compared with the costs in the rill stopes. All the ore cannot be carried to the chute, in falling from place, unless the chutes are very close together or the distance from the back of the stope to the top of the chute is sufficient to slide all the ore broken. The number of chutes is limited by the tendency of the ground to cave; if carried too close together the stope is weakened. Small stopes, or blocks,

Jan 1, 1923

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By Blamey Stevens

(Butte Meeting, August, 1913.) THE following paper aims to make a full explanation of the phenomena of rock jointing: It may be unnecessary to give any general description of what are termed joints in rocks, but Professor Geikie's clear and faithful generalizations are so illuminating that no excuse need be given. for introducing the following and other quotations into this paper "Almost all rocks are traversed by vertical or highly inclined divisional planes termed joints. These have been regarded as due in some way to contraction during consolidation (fissures of retreat) ; and this is no doubt their origin in innumerable cases. But, on the other hand, their frequent regularity and persistence across materials of very varying texture suggest rather the effects of internal pressure and movement within the crust." 1 General Explanation.-In this paper I show that the phenomenon of jointing may to a large extent be accounted for by the pressure of the water contained in the pores of the rock. To understand this, each constituent grain of the porous material may be considered to be separately compressed by the water which submerges it. As all the grains are held together by cohesion, the whole mass is contracted by this pressure. In an indefinitely extended mass, however, there must be a limit to the possible contraction without rupture, because the material is restrained from moving freely by other material around its boundaries. The mass of rock thus becomes broken up. This phenomenon is known as '" shrinkage " jointing. The term " shrinkage " is an apt one, if it be understood that the shrinkage of the rock is often brought about by change of stress. Stresses.-The compressive component of water stress on the individual grains is evidently the same as that of the water with which it is in contact. This is so, both before and after the jointing is formed. Before the jointing is formed there is, however, a tensile stress in one or more directions which keeps the material from contracting. .

Jan 7, 1913

By I. Iwasaki, K. Adam

During wet grinding the electrochemical interactions between grinding media and pyrrhotite not only accelerate media wear but also adversely affect the suede mineral's flotation response. A possible remedy of the adverse effects on floatability by applying electric potentials to grinding media was examined. Flotation tests, Auger electron spectroscopy, and X-ray photoelectron spectroscopy were used to study pyrrhotite surface properties. Although anodic polarization of the steels increased metal loss by corrosion, pyrrhotite flotation recovery was increased. In the case of cathodic polarization, flotation recovery improved somewhat in a narrow range of overpotentials. At higher cathodic overpotentials, however, mineral floatability was progressively lowered although the corrosive wear of grinding media might be reduced.

Jan 1, 1985

By M. S. Miller, R. H. Mullennex

The growing need for development of thinner, more erratic, or less easily-mined coals in the Appalachian region requires more detailed analysis of conditions and factors controlling or influencing deposition of these seams. The value of depositional modeling in evaluation of such coals has been demonstrated in many recent studies and publications. Such modeling may be applied to allow prediction of factors varying from regional coal occurrence trends to very local areas of seam discontinuities but requires thorough we of all available types of data. Mapping of general depositional energies, as derived from basic geophysical and driller's logs, can be applied to specific prediction of coal occurrence, and projection of particular facies as representative of unique depositional features or conditions may prove more applicable to coal character delineation than mapping of the coal itself. Depositional modeling and trend projection may also serve as a guideline in reserve classification and confidence levels in reserve estimations.

Jan 1, 1983

By E. E. McGinley, L. D. Woodworth

DURING the past several years has occurred what, in the light of future events, may aptly be called the rebirth of the acid Bessemer process. The increased attention given to the technical and metallurgical details involved in the production of Bessemer steels has been exceedingly fruitful, and indicates the value of further research and plant development. In times like these, when the steelmaking capacity of the country is operating at a maximum rate to supply materials for the national defense program, it behooves the Bessemer operator not only to strive to increase production but also to maintain a high standard of quality. The opportunity is ripe, it seems, to recapture fields of application once considered the birthright of Bessemer steels, but supplanted in recent years by open-hearth grades. This involves cooperation all along the line, from the blast-furnace operator to the metallurgical contact man and salesman. The metallurgist's part in this program involves the study of the many steelmaking variables and the measures necessary to ensure constancy in the finished product. It is equally important that the knowledge gained be disseminated widely to ensure the wise application of Bessemer steels in new fields. The merits of Bessemer steels for certain work are well understood. The combined properties of good weldability, machinability and stiffness are unique and fully explain the strong entrenchment of Bessemer steels in certain fields; e.g., skelp for small butt-weld conduit, screw steels machined in automatics, and certain tinplate applications. It is the purpose of this paper to review briefly the physical properties of Bessemer steels as contrasted with those of open-hearth steels, to discuss control measures now employed in the making of Bessemer steel, and to consider the advantages derived from the use of these control measures. PHYSICAL PROPERTIES OF BESSEMER VS. OPEN-HEARTH STEELS A number of Bessemer blows and open-hearth heats were tested for tensile strength, ductility and impact strength. The steels selected were confined to sections from ¾ to 1 ¼ -in. diameter, inclusive, and the carbon content covered was sufficient to yield parallel ranges in ultimate strength for the comparison. Other factors, such as the extent of deoxidation of the steel and variations in chemistry, were kept as comparable as possible. Machined test pieces of 0.505-in. diameter with an 8-in. gauge length were pulled. The data obtained are shown in Fig. 1, ultimate strength being plotted against carbon content. This graph shows that for equivalent carbon content Bessemer steel has on the average 15,000 lb. per sq. in. greater tensile strength and, conversely, for the same tensile strength Bessemer steel requires on the average 0.14 per cent less carbon. The average phosphorus content for the Bessemer steel was 0.090 per cent,

Jan 1, 1941

By Kevin B. Doyle

The export markets are of great importance to western Canadian sulfur producers, taking 70% of total products sold. For these marketers, Canadians have always supplied solid sulfur, first in a crushed, bulk form, and for the past several years in a slate form. Both form have proved inadequate for safety and environmental reasons, and new forms and processes have been developed by interested suppliers. The Canadian industry has evaluated these processes and ha made a determination of a desired Product specfication that will greatly enhance the acceptability of future Products from safety, environmental, and customer points of view. The new processed methods and resulting product characteristics are reviewed.

Jan 1, 1980

By Keith G. Papke

Jan 1, 1985

By E. D. Meier

SINCE Dr. St. Albans, in 1840, began to build successful watertube boilers—some of which are still in use—the gravity-return water-tube boiler has been built in many forms, more or less familiar to all who have studied the subject. A single firm in this country published recently a pamphlet, showing some fourteen modifications of their boiler during about as many years. There were hence available to the thoughtful inventor of this latest form many data showing what forms were to be avoided, as well as the effects sure to follow certain erroneous methods of construction in the older types. Improvements in materials, tools, and workmanship made possible, forms formerly prohibited by excessive cost. The first Heine boiler was built by A. Borsig & Co., for the

Jan 1, 1886

By M. R. Daly

In a recent paper on Geologic Structure in the Cushing Oil and Gas Field, Oklahoma,l Carl H. Beal has pointed out some interesting peculiarities in the distribution of the hydrocarbons and the disposition of the "planes of separation" between the oil and the water (water surfaces). Mr. Beal's principal statements on this matter may be, summarized as follows: I. In every important dome in the Cushing field, which has not been complicated by folding along its sides, the oil area in each of the three oil and gas sands studied (Layton, Wheeler and Bartlesville) apparently extends farther down on the long axis of the dome than it does on the steeper sides. The result is that water is encountered at a lesser depth below datum plane (sea level) on the steep sides of the dome than along the ends where the beds are more gently tilted. II. In the Layton and Wheeler sands (to which the detailed study of water surfaces has been almost entirely confined), the water surface is not level, but dips at a gentle angle (maximum dip 100 ft. to the mile) away from the center of the structure. This inclination existed prior to development, and it has been rendered greater by the rapid extraction of oil and gas. To these statements, the following observations are added: The dominant structural feature in the Cushing field is a broad north and south anticlinal fold, along whose axis the domes are distributed (Fig. 1, p. 897), and the more important oil areas in each sand lie on the west side of the field. In many cases, gas has been forced to the east side of the structure, sometimes as far as the water line. From the distribution of the oil and gas bodies, Mr. Beal concludes that the hydrocarbons in this region probably migrated from the west or the northwest. It seems as if these different features would be susceptible of some

Jan 1, 1918

Some 30,000 people attended from 50 countries. Over 600 exhibitors brought with them a billion dollars of mining technology, ranging from tiny high-precision valves to giant off-highway trucks. Twenty-seven technical sessions with more than 100 papers covered the gamut of mining advancements. It was a grand display of industry vitality just to power the huge exhibit area required 22,000 kW of electricity each day, and support services for the exposition created more than 1000 daily jobs for local labor. Unquestionably, Las Vagas knew the miners were in town for the American Mining Congress' International Mlining Show.

Jan 11, 1978

By O. Knacke, H. Barth, W. Gans

The rate of cementation was determined by measuring the time required to consume a thin electrodeposited layer of the cementating metal as indicated by a sudden increase in the electrical potential. The diffusion controlled cathodic current of the more positive metal was measured by conventional electrochemical techniques and compared with the cementation rate under the same experimental conditions. Micrographs revealed different types of crystal growth in different surroundings e.g. dendritic growth in acidic solutions.

Jan 1, 1973

By Robert Spring

The kinetic ball mill model accurately simulated two industrial mills over their entire operating range. This was confirmed using 100 sets of grinding circuit data. Contrary to initial expectations, the absolute grinding rate function was not correlated to the mill feedrate, feed size distribution or pulp density. Instead, a single rate function provided excellent simulation over the entire operating range. Statistical analysis showed that the model explained all significant variation in measured product size distributions.

Jan 1, 1984

By L. W. Pommier, F. B. Brien

In order to determine the ability of cyclones with heavy liquids to preconcentrate ores, a study of the variables which govern the behavior of the solids in the apparatus has been made. To assess the test results, the cyclone products have been examined by sink and float analysis at various specific gravities and the results examined by the Tromp type curves. This analysis has allowed the development of the SG50 unit — specific gravity of the particles, 50% of which goes to overflow and 50% to underflow — rather than the d50 (size) unit used in classification. The results of tests using the closed circuit laboratory cyclone assembly with dry feed samples show the validity of the equation: Results indicate that 83% of the feed may be rejected containing 0.1% Sn. Recoveries in excess of 80% are obtained at a grade of 1.7% Sn from feed samples assaying 0.4% Sn. As water-wetted particles may comprise the plant feed, separate tests were performed to examine the effect of water in the feed material. Results indicate that the entrained water divides into two fractions, one of which is defined as free water and is rejected to the cyclone overflow, the other fraction forms an adhering envelope of water on the ore particles, altering their effective size and density. The presence of the envelope of water on the particles increases the ease with which the heavy liquid can be recovered but decreases the concentration efficiency. It is proposed that the detrimental effect of water in the cyclone may be minimized by using a higher feed pressure cx by the use of surfactants. A higher feed pressure, by increasing shearing forces, would reduce the thickness of the water envelope on the particles. The surfactants, by adsorbing on the particles and producing hydrophobic surfaces, would displace the water envelope and allow the heavy liquids to wet the modified surfaces. In both cases a larger proportion of the entrained water would report to the cyclone overflow as free water, thus minimizing the detrimental effect of water on the concentration efficiency. Additional testing is required on the behavior of water in the system; to determine the optimum particle size for the preconcentration step; to determine final particle size for optimum liberation and subsequent treatment of the products in present plants and to determine the efficiency with which the heavy liquids may be recovered from the products. For many years tin mining in Bolivia has been very important to the economy of the country. In recent years the tin content of the ore reserves has been steadily declining. An improvement in concentration efficiency is highly desirable in order to increase productivity from the present low-grade mine ores and to allow retreatment of the many millions of tons of tailings from past operations, some of which equal the present ores in tin content. A number of investigators have been or at present are working on various phases or approaches to the problem. Gravity concentration is used in Boliva today as it has been in the past to recover tin minerals from the mine ores. An objective of the work reported in this paper was to examine the possibility of treating present low-grade ores and sink-float tailings products, at a relatively coarse grind, in order to produce a higher grade feed material to the present gravity concentration plants. The possibility of using heavy liquids in cyclones to treat large tonnages with a minimum of outlay presented itself and is of particular interest since heavy liquids are now available with physical properties which lend themselves to simple recovery methods. Tests were carried out using a closed circuit laboratory cyclone assembly. It was desirable to determine the value of using the SG50 concept and Tromp curves for analyzing the cyclone concentration test results. The effects of entrained water on the behavior of ore particles in the cyclone were examined. THEORY AND TECHNICAL CONSIDERATIONS Three types of heavy liquids are available: aqueous solutions of very soluble salts, salts or mixtures of

Jan 1, 1964

By C. Wegert, L. E. Paulson, S. A. Cooley, R. C. Ellman

The Grand Forks Energy Research Center (GFERC) and Commonwealth Edison of Chicago jointly conducted tests in which 400 tons of subbituminous and lignite coals were dried in a commercial scale dryer, oil sprayed and cooled, then shipped from Pekin, Ill., to Grand Forks, N.D., and stockpiled. Cars containing raw coal and dried coal which had not been sprayed, were also transported for comparative purposes. During transit, observations were made of changes in the coals' moisture content, dust losses, and temperature.

Jan 1, 1977

By William M. Brewer

Since writing the paper on this subject,* which was presented at the Atlanta meeting in October, 1895, I have had opportunity for more thorough investigations in several localities, and venture, therefore, to offer some additional facts and opinions. The literature of the subject is not very abundant. Dr. George F. Becker's " Reconnaissance of the Gold-Fields of the Southern Appalachians "† contains some brief notes relating to Georgia, but does not include Alabama. The paper of Messrs. Nitze and Wilkens on " Gold-Mining in the Southern Appala-

Jan 1, 1897