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By J. Kostuik, M. J. de Bastiani

Located in the Elliot Lake mining district of Canada, the Denison mine lies between Sudbury to the east and Sault Ste. Marie to the west. This area is overlain by Huronian sediments filling a basin in the Keewatin basement complex. Subsequent folding of these sediments developed a synclinal structure, known locally as the Quirke Lake Trough, which forms the northern portion of the now famous big Z structure, the geological trademark of this region. Plunging to the west and extending nine miles from the granite outcropping in the north to the Keewatin lava contact to the south, the syncline has dips of 10° to 45° on the north limb and 10° to 20° on the south limb.

Jan 12, 1960

By J. F. Poland

Location of Basins and Geologic Features of Occurrence: The major ground-water resources of California occur and are stored in the many large alluvium-filled valleys of the state. The deposits of Quaternary age which hold the ground water commonly are called valley fill. Meinzerl has discussed the characteristic features of this fill and showed that for western United States its distribution is concentrated chiefly in western Utah, Nevada, central Oregon, California, southern Arizona, and central New Mexico (see fig. 1). For California the principal ground-water basins are shown on fig. 2. Except for the Mojave Desert and the "Basin and Range" Valleys (fig. 2, Nos. 23 and 23A), which are not shown individually, the basins outlined on fig. 2 contain all the large reservoirs in valley fill within the state. The Sacramento and San Joaquin Valleys (fig. 2) together comprise the Great Central Valley of California, which has an area of about 18,000 square miles. The Sacramento Valley contains about 5000 square miles of valley lands, approxi- mately equal to the area of Connecticut; the San Joaquin Valley embraces about 13,000 square miles, equivalent to the areas of Massachusetts and Connecticut combined. Both the area and the ground-water use in the Central Valley is far greater than for all the other shaded alluvial basins of fig. 2 added together. Except for the Sacramento Valley, all the major ground-water basins of the state are south of Sacramento. The valley fill in these ground-water basins has been transported by streams and most of it has been laid down in alluvial fans, producing the typical alluvial-fan structure of aquifers radiating as tongues from a prismatic pile of coarse, permeable debris at the canyon mouths; these aquifers are encased down slope by poorly sorted materials consisting of a clayey matrix with imbedded pebbles and boulders, interbedded with flood-plain silt. In the deeper inland structural valleys, the alluvial deposits interfinger at depth in the valley bottoms with horizontally bedded lacustrine sand, silt, and clay. The San Joaquin Valley is a good example. In the coastal valleys exposed to the sea, such as the Los Angeles coastal plain, the alluvial fans pass coastward into lagoonal and thence to shallow marine deposits. In all these basins, the gravel is coarsest and most permeable near the apexes of the fans, but tongues or lenses of clean gravel and sand may extend for many miles down the fan. Thus, these water-bearing structures are in a position to absorb stream flow near the canyon mouths and to transmit it through the buried radiating conduits.

Jan 1, 1951

By J. F. Poland

Location of Basins and Geologic Features of Occurrence: The major ground-water resources of California occur and are stored in the many large alluvium-filled valleys of the state. The deposits of Quaternary age which hold the ground water commonly are called valley fill. Meinzerl has discussed the characteristic features of this fill and showed that for western United States its distribution is concentrated chiefly in western Utah, Nevada, central Oregon, California, southern Arizona, and central New Mexico (see fig. 1). For California the principal ground-water basins are shown on fig. 2. Except for the Mojave Desert and the "Basin and Range" Valleys (fig. 2, Nos. 23 and 23A), which are not shown individually, the basins outlined on fig. 2 contain all the large reservoirs in valley fill within the state. The Sacramento and San Joaquin Valleys (fig. 2) together comprise the Great Central Valley of California, which has an area of about 18,000 square miles. The Sacramento Valley contains about 5000 square miles of valley lands, approxi- mately equal to the area of Connecticut; the San Joaquin Valley embraces about 13,000 square miles, equivalent to the areas of Massachusetts and Connecticut combined. Both the area and the ground-water use in the Central Valley is far greater than for all the other shaded alluvial basins of fig. 2 added together. Except for the Sacramento Valley, all the major ground-water basins of the state are south of Sacramento. The valley fill in these ground-water basins has been transported by streams and most of it has been laid down in alluvial fans, producing the typical alluvial-fan structure of aquifers radiating as tongues from a prismatic pile of coarse, permeable debris at the canyon mouths; these aquifers are encased down slope by poorly sorted materials consisting of a clayey matrix with imbedded pebbles and boulders, interbedded with flood-plain silt. In the deeper inland structural valleys, the alluvial deposits interfinger at depth in the valley bottoms with horizontally bedded lacustrine sand, silt, and clay. The San Joaquin Valley is a good example. In the coastal valleys exposed to the sea, such as the Los Angeles coastal plain, the alluvial fans pass coastward into lagoonal and thence to shallow marine deposits. In all these basins, the gravel is coarsest and most permeable near the apexes of the fans, but tongues or lenses of clean gravel and sand may extend for many miles down the fan. Thus, these water-bearing structures are in a position to absorb stream flow near the canyon mouths and to transmit it through the buried radiating conduits.

Jan 1, 1951

By Thomas Elliott

It is now clearly evident that 1963 is witnessing a record year for prospecting and exploration in British Columbia and Yukon Territory, as substantiated by the following facts: ? About 60 large mining companies and many smaller ones are active in the search for new ore deposits. ? More than 500 prospectors and exploration geologists have been in the field this season. ? It is estimated that $10 million has been spent this year on the search for new ore deposits. ? Claim staking in B. C. to July 31, totalling 14,250, is up 2000 over 1962. ? Helicopter companies report a record total of 25 aircraft chartered to mining groups, with additional numbers used in oil-gas exploration. ? Leading assayers state that business with the industry is double that of the previous year. ? Diamond drillers report an increase of approximately 25% in the amount of contracts now on hand. The question is often asked, "Why is mining so active in Canada's Pacific Northwest?"

Jan 12, 1963

By C. A. Stetefeldt

The iron-ore deposits worked by this Company occur in lenses 200 to 1000 feet long and 5 to 80 feet wide, and stand at an angle of from 65' to 75, with a vertical height of 250 to 500 feet, other lenses occurring below. A number of the deposits were first worked as open pits, which in some cases were carried to the depth of 150 feet, when, owing to the weakness of the walls, underground mining was adopted. While the ore was being removed from the open pit, shafts mere in several instances sunk into the foot-wall, the intention being to mine the ore with breast-stopes of an approximate height of 20 feet, followed by underhand-stopes of the same height, leaving floors between of the necessary thickness to support the walls, or to effect the removal of the ore by means of what is known as the stall system. As the work progressed, however, it was found that the walls (which are of chlorite) were too weak to permit the working of breast-stopes 20 feet high, there being frequent heavy falls of ground from the hanging and, in some instances, from the foot. The plan of following breast-stopes with underhand-stopes was, therefore, abandoned. By working breast-stopes only, but little more than one-half of the ore could be removed, and that only at an excessive cost, the ore being so hard that power-drills with 3 1/8-inch pistons . and 6-inch stroke, working under 60 pounds of compressed air, are able to drill but 8 feet in ten hours as a yearly average, while from 6 inches to 2 feet is a common result of ten hours' drilling. It was plain that some other system of mining must be adopted, and it was proposed to sink by levels of 75 feet, carrying in the cross-cuts from the shafts and working out the ore each way from the shaft from foot to hanging, and from 15 to 20 feet in height. When this has been done, drift-sets consisting of caps and legs are set up the whole length of the opening and connected with the cross-cut; the necessary openings for ladder-ways and chutes or mills are timbered from

Jan 1, 1893

By D. S. Choi, D. L. McCain

A method of estimating the chain pillar size required to support the wide working faces of longwall or shortwall coal mining is presented. The pillar size is determined through the use of coal strength, entry geometry, panel width, and overburden pressure. Design analysis is focused on a three-entry development system in the Pittsburgh Seam but can be used for other development plans and other seams by substitution of the proper geometric and strength data.

Jan 1, 1981

By Robert R. French

Beneficiation of low-grade nonmetalliferous mineral deposits by electronic color sorting is currently undergoing a period of rapid development. For those deposits in which the specific gravity, strength, soundness, and electric and magnetic properties of the ore and gangue are similar, the possibility of separation by color, reflectivity, or fluorescence offers hope for the utilization of much poorer quality ore than previously thought possible. For those ores in which consistency of color is required or for which a dry process is preferred, the electronic color sorter is particularly adaptable. Tests performed late in 1966 and reported herein are the result of continued interest in the dry beneficiation of gypsum and the development of higher capacity color-sorting machines. These tests should be considered only as an indication of the efficiency of the process and only with respect to the low-grade, untreated ore tested. Although most of the ore mined in Martin County, Ind., is of acceptable quality, a few thin carbonate strata and irregular nodules of dolomite and limestone are present within the main evaporite bed (Fig. 1).

Jan 1, 1969

By S. D. Warriner

ONE of the most, if not the most, difficult of the problems associated with the mining and distri-bution of anthracite is to get clearly into the minds of the consumers and of the editorial and reporto-rial writers in the public press the relationship between the cost of production, the price at the mine and the price that the consumer pays for his coal, delivered in the cellar. One sees in his morning paper a statement to the .effect that out of the $12 or $13 per ton that his dealer charges him, the miner gets only $2 or even less, and he asks himself or the paper asks for him, "Who gets the difference?" Not infrequently the amount that the miner gets is given for run-of-mine bituminous coal while the price quoted is that of prepared anthra-cite, egg, stove, or chestnut, as the case may be, delivered to the bin of the consumer; and the difference stands out with all the more prominence; and it sticks in his memory with a persistency which no manner of explanation is able to efface. A recently published official report for the year. 1921 of one of the State Bureaus stated the total production of anthracite in net tons of 2000 lb., and included not only the output recovered by washeries from old culm banks and by dredges from creek and river channels, but also the colliery consumption;, i.e., the coal used by the companies in the operation of the collieries. This resulted in a total production of 90,509,075 net tons for which the labor cost was stated at $283,961,300, or $3.13 per net ton, equivalent to $3.50 per gross ton. This figure of $3.50 was immediately seized upon as representing the labor cost per gross ton of fresh-mined commercial coal. It is almost a hopeless task to try to correct such errors of statement in the first place, and, in the second place, to correct the misapprehen-sions that arise from them. As a matter of fact, the report was misleading in regard to the labor cost of the entire production shipped, as the coal used for colliery consumption should be deducted. Correctly stated, the report should have shown, calculated in terms of gross tons of 2240 lb., the legal weight in Pennsylvania, as follows: Gross production 80,607,359 tons Less colliery fuel 8,757,137 tons Commercial production 71,850,222 tons Total labor bill $283,961,300.00 Labor cost per ton...: 8 3.952 The foregoing figures" of tonnage are obtained from the State Department of Mines. Even these include the quantity recovered from calm, banks by washeries, and consequently the $3.95 arrived at as the average labor cost is somewhat less than that actually involved in the mining of fresh-mined coal. Unfortunately, the figures of washery production and the cost of producing it are not available.

Jan 7, 1922

By Tony B. Szwilski

The paper outlines the parameters required for an estimation of the size of chain pillars for longwall coal mining. The principal parameters are: stress distribution around the edge of the extracted coal panel, strength properties and geometry of the pillars. The principal thrust of the paper is directed toward correlating the chain pillar strength to the magnitude and direction of strata load expected to act on the pillars. A brief review of the pillar strength formulae pertinent to chain pillar stability is given. The chain pillars are categorized as 'yielding' or 'stiff', depending on the duty required of them. The coal measure strata is classified as 'soft' or 'hard' conditions, the degree of competence of the roof and floor greatly influencing the caving characteristics of the gob and the stability of the roof chain pillars.

Jan 1, 1983

By George L. Adams

There is an extensive literature relative to gas, oil, and the more solid hydrocarbons; but when it has all been digested and summarized, the resulting information is far from being satisfactory. The problems which have received the most attention are naturally those which relate to the successful exploitation and utilization of the deposits. From what is known of their geologic occurrence, theories have been formulated and rules laid down for the guidance of prospectors. The technology of the products has been made the subject of special chemical investigations. The geologic interrelation of the hydrocarbons, however, has not been considered in a broad manner, and the principles which have governed the origin of the economic bodies have been but imperfectly expressed. On the other hand, metalliferous deposits have been widely

Jan 1, 1903

Jan 1, 1927

Jan 1, 1889

By R. M. Asimow

The magnetization curves for specimezs containing superparmagnetic particles are considered. It is shown that the curves may differ from a Langezlin function because of particle anisotropy, particle interactions , and the presence of a size distribution of particles. The various causes of particle anisotropy are considered, and magnetization curves for specimens with cubic and uniaxial anisotropy are presented. The magnetization curves depend on the In the particle-size below 100A it is quite difficult to obtain quantitative size and shape information on precipitate particles. All of the available techniques have their limitations; however, quantitative applications of superparmagnetism have not been developed as fully as possible due to a lack of analysis of the complete magnetization curve. With such an analysis it appears possible to obtain from a single magnetization curve of a specimen containing randomly oriented SPM (superparamagnetic) particles a quantitative volume distribution and a description of the particle shape.* sign of the anisotropy energy and in the case of a negative uniaxial anisotropy (prolate spheroids) the resulting curves may differ very considerbly from a Langevin function. A graphical technique is developed whereby it is possible to separate out the effects of the size distribution, the particle interaction, and the particle anisotropy. Thus it is possible to obtain quantitative values for both the particle shape and the size distribution. Langevin function2 I/Is = ctnh(HVIs/k T)-k T/HVIS [ 1 ] where I, is the saturation magnetization, V the particle volume, H the applied field, T the particle temperature, k the Boltzmann constant, and I the magnetization in the direction of the applied field. Such behavior is rarely if ever observed in actual systems. One explanation for this is that the SPM particles do possess a significant anisotropy and thus the magnetic energy associated with the particle depends on the direction of magnetization in a more complicated way than for the isotropic particle of Eq. [I]. In general the magnetic energy E of an isolated SPM particle embedded in a nonmagnetic matrix can be expressed as a sum of several contributions: E = Eh + Ed+Es+Ec [2] where Eh is the external field energy, Ed the energy resulting from the particle demagnetizing factor, E, the magnetostriction energy, and E, the crystalline anisotropy energy. In each term we consider only the part dependent on the orientation of the magnetization vector and the particle axes as the shape of the magnetization curve will be determined by these contributions. Following the same procedure used in deriving the Langevin function,

Jan 1, 1965

Jan 1, 1909

By A. W. Hahn

Although enormous tonnages of sulfide lead and lead-silver ores are treated by flotation, the products of flotation mills treating oxidized ores of lead and silver are almost negligible. However, each year there is a slight tonnage increase, and it is possible, should a sizeable oxidized lead ore tonnage be developed, the field will become an important one. In 1918, the Shattuck-Arizona Copper Co. erected a plant at Warren, Ariz., for the treatment of oxidized gold-silver-lead ores. Glenn L. Allen worked out the metallurgy for this plant, employing tabling, followed by flotation concentration. The Shattuck-Denn Mining Co., successors to the Shattuck-Arizona Co., now employs straight flotation in the same mill, treating about 200 tons daily. In 1922, the San Diego mill of the American Smelting & Refining Co., in the Parral district, Chihuahua, Mexico, started operating a 600-ton mill on a mixed sulfide-oxide gold-silver-lead ore. Tables and flotation are used here, and about 200 tons of oxidized slime treated by flotation. In 1923, the Eureka Holly Mining Co. started operating a 90-ton mill at Eureka, Nev., using tables and flotation, on an oxidized lead-silver ore. In 1924, the American Smelting & Refining Co, started a 250-ton mill in northern Peru, to treat a high-grade oxidized gold-silver ore with very low lead values. In 1924, the Grand Central Mining Co. started a 200-ton flotation mill on old oxidized gold-silver-lead tailings at Fairbank, Ariz. In 1926, the Compania Internacional Minera, a subsidiary of the Union Corporation of London, started a 250-ton flotation mill on oxidized lead-silver ores, and during April of this year, the Phelps-Dodge Corpn. will probably start a 250-ton mill at Warren, Ariz., on oxidized lead-silver ores, similar to those being treated in the Shattuck-Denn mill. Methods All of these mills use sodium sulfide for filming the oxide minerals, with the possible exception of the North Peru mill, and all use oily reagents. The total tonnage of strictly oxidized lead-silver ores concentrated by flotation is probably less than 2500 tons daily. The following figures give the various types of ores being treated:

Jan 1, 1928

By C. W. Greenhalgh

THE accuracy with which grade and tonnage of a mineral deposit may be calculated from diamond-drill assays is dependent upon: (I) the reliability of core and sludge samples and (2) the validity of methods used in grade and tonnage calculations. Limitations imposed by drilling and sampling preclude elaborate methods of computation. In adjusting core and sludge assays the importance assigned the core and sludge samples may be based on either the respective weights recovered or the respective volumes represented by each. Only when core and sludge recoveries are poor do discrepancies between adjusted assays calculated by these two methods develop, in which case the "weight" method gives emphasis to the core sample and the "volume" method gives emphasis to the sludge sample. When the sludge suffers classification within the drill hole, the weight method gives more accurate results; when the sludge sample suffers no classification, the volume method gives more accurate results. Any drilling campaign offers a compromise between desired accuracy on the one hand and expediency, limited time, money, or facilities on the other. Since the determination of grade and tonnage is an approximation at best, a proper basis for their computation requires a knowledge of the limitations imposed by all phases of work upon which the calculations depend. A rational basis for coordination of these phases of work requires determination of: (I) the degree of accuracy desired, as a function of the object of examination; (2) the degree of accuracy attainable, as a function of the limitations imposed in the drilling and sampling program; and (3) the validity of method to be used in calculating grade and tonnage from core and sludge assays. This paper presents results of a comparison made between two commonly used methods of adjusting core and sludge assays, the assumptions upon which these methods are based, and the interrelationship of the phases of work upon which these calculations are dependent. OBJECT OF EXAMINATION The purpose of prospecting any mineral deposit is to secure sufficient samples from which ore grade and tonnage may be calculated. Any mine examination, preliminary or formal, has a definite purpose, although its scope may be limited or broad. The type of examination to be conducted is dependent, therefore, upon the reason for instituting the investigation; its scope and thoroughness are directly related to economic considerations. It is highly desirable, therefore, that before starting a drilling campaign a complete understanding of the extent of examination and desired degree of accuracy be determined as a means of simplifying the drilling program and unifying the relative degree of

Jan 1, 1946

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the dose association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and' 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill 10,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. 1) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By A. C. England, William F. Haney, Paul B. Crawford, Bobby L. Landrum

The creation of unsymmetrical vertical or horizontal fractures in a producing well will result in a unique distortion of the normal shut-in pressures existing at the four surrounding wells. An electrical model study has been made to show how the shut-in pressures at the wells surrounding the fractured producing well may he used to estimate the size and orientation of vertical or horizontal fractures. The pressure differences at the four surrounding wells are recorded during production from the fractured well. The resulting pressure dato are matched from a table of several hundred possible fracture combinations. The table refers to the figure showing the effective size and orientation of the commercially created fracture. INTRODUCTION The creation of commercial fractures about oil and gas wells has led to considerable speculation as to the size and shape of the fractures. It is important that the size and shape be known in certain instances, especially when pressure maintenance programs are to be initiated. It was shown that if knowledge of a vertical fracture could be found that such information may be utilized in planning a waterflooding or gas cycling program and that in some cases the fractures might be utilized very effectively.' Other work showed that it may be desirable to create "pancake" fractures in certain reservoirs but that they should exist in a desirable range.' If the fractures are too large, they should be plugged prior to initiating a pressure maintenance program. If an irregular shaped "pancake" fracture is created, it is desirable to know the orientation of this fracture because similar but unreported studies have shown that "pancake" fractures may or may not substantially affect the sweep efficiencies and performance of the flooding or cycling program. Consequently, it may be desirable in some instances to estimate the effective size and orientation of the fracture. ANALOG An electrical model was used in making this study. The electrical model consisted of a circular reservoir approximately 8 in. in diameter, 1 in. thick and was filled with a dilute salt solution. A copper band formed the periphery of the reservoir. A vertical fracture was simulated by a copper strip soldered to the wire well located in the center of the reservoir. Vertical copper strips of various lengths and orientation were soldered to the wire well to represent vertical fractures in the reservoir. In studies reported here on vertical fractures it was arbitrarily assumed that two fractures originated at the well and extended out into the reservoir. An electrical potential was imposed between the wire well and the periphery of the tank. Plots of the iso-potentials are shown in Figs. 1 through 9. This model provides the steady-state isopotentials existing in a uniform, homogeneous circular reservoir having a vertically fractured well at its center. The permeability of the fracture is much greater than the permeability of the matrix proper. The effects of gravity are neglected. In preparing Tables 1 and 2, four symmetrically located unfractured wells were assumed to exist around

Jan 1, 1956

By J. W. Downey, A. E. Dwight, R. A. Conner

THE MgCu2 type Laves phase has been observed in the thorium-osmium and thorium-iridium systems. Two alloys corresponding to ThOs2 and ThIr, were prepared by arc melting a 5-g charge in a water-cooled copper. crucible. Both alloys were very brittle. X-ray diffraction patterns were obtained using filtered Cu and Co radiation with a Straumanis-type DeBye-Scherrer camera. The a,, values as shown in Table I were obtained by a Nelson-Riley extrapolation. The structure belongs to space group Oh7-Fd3m. Relative intensities for each of these compounds were calculated on the assumption that 8 thorium atoms occupy (a) positions and 16 osmium or irid-ium atoms occupy (d) positions. The calculated intensities are the product of the Lorentz-polarization factor, the multiplicity factor and F', the structure factor, neglecting the absorption and temperature

Jan 1, 1959

By Robert Franke

(Butte Meeting, August, 1913.) IN view of the prominence which the conical mill has attained in the fine-crushing field .within the few years since its introduction, the following comparison with its more mature forerunner, the Chilean mill, based on extensive tests, is submitted in the interest of the milling profession. Soon after designing the concentrating plant of the Miami Copper Co. in 1909, the Hardinge conical mill made its appearance in the milling forum. . Its possibility as a suitable crushing device for the plant was well recognized, but in view of the lack of commercial demonstration, at the time, as to capacity, efficiency for desired product, and the still more uncertain factors of cost of maintenance and power consumption, it was deemed that the immediate adoption of this machine throughout the plant would be a hazardous undertaking. For these reasons it was decided to equip the majority of the immediately required units of the plant with Chilean mills, the fine-crushing proficiencies of which were better known, and one section with Hardinge mills, to serve as a test unit for the guidance of future installations and replacements. Thereby, after 1.5 years' operation with both types of mills, a thorough test as to metallurgical efficiency and cost economy has been obtained. The conical mill used in these tests is the 8-ft. Hardinge pebble mill, having a cylinder 22 in. in length. The cylindrical portion of this mill is lined with cast-iron liner plates, and the conical extensions with silex bricks bound together by cement. Each. liner plate carries a projecting lifter, the function of which is to increase the height of drop of the lifted material. Danish No. 5 pebbles, obtained from the coast deposits of Denmark, are used for the grinding charge. The Chilean mill used is a fast-running, 3-roller, 6-ft. Saturn mill, with screens of 0.037-in. opening. The feed to these mills is the oversize of Callow screens having 0.029-in. openings, which. follow rolls crushing to 0.5 in.

Jan 7, 1913