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By W. F. Holbrook, C. E. Wood, E. P. Barrett

The high-frequency laboratory induction furnace with a rotating liquid crucible enables research workers to conduct certain investigations heretofore very difficult or impossible to realize because vessels are not obtainable that are capable of resisting the physical conditions or the chemical actions to which they are subjected. Applications for patents covering commercial uses of centrifugal liquid crucibles were filed in the United States in 1919,' in France in 1925, in the United States and Germany in 1926,2 and in France in 1927.3 In 1928, Schuette and Maier,4 of the Bureau of Mines, reported the construction of a whirling table beneath a high-frequency induction furnace. Shape and Size of Centrifugal Liquid Crucibles When a vessel containing a liquid is rotated, the inner surface of the liquid takes the shape of a paraboloid under the simultaneous action of centrifugal force and gravity. The internal dimensions of the paraboloid rotating liquid crucible are governed by the diameter of the container and its speed of rotation. Melted Wood's metal was poured into cylinders 2 1/2 in. in diameter, which were rotated at 180, 300, and 468 r.p.m. When the metal was solid, a slurry of plaster of Paris was added, and rotation continued until the plaster had set. Fig. 1, a photograph of sections of the cylinders, shows the effect of speed of rotation of the containers upon the shape of the inner surface of the rotating liquid crucibles. Types of Furnaces Three types of high-frequency induction laboratory furnaces using liquid-metal crucibles were designed and constructed by the Blast

Jan 1, 1939

Jan 1, 1920

I. PRESIDING OFFICERS. At all Business meetings of the Institute the President, or, in his absence, the Vice-President, or, in the absence of both of them, any other member of the Board of Directors to be chosen by the meeting, shall preside. At all other meetings of the Institute the President of the Council or, in his absence, one of the Vice-Presidents, if present, shall preside II. ORDER OF BUSINESS. At each Business meeting of the Institute the order of business shall be as follows 1. Reading of minutes of preceding meeting. 2 Report of the President. 3. Report of the Treasurer. 4. Report of the Secretary. 5. Election of Directors. 6. Election of Members of the Council. 7. Reports of Standing Committees. 8. Reports of Special Committees. 9. Special Orders 10. Miscellaneous business. This order of business may be changed by a vote of a majority of the Members and Associates present in person or by proxy. The usual parliamentary rules shall govern all meetings of the Institute except in cases otherwise provided by the Constitution or the By-Laws. At all sessions of the Institute other than business meetings, the order of proceedings and the time of adjournment, shall rest in the discretion of the presiding officer

Jan 1, 1917

I PRESIDING OFFICER At all business meetings of the Institute the President, or, in his absence, the First Vice-President, or, in the absence of both of them, any other Vice¬ President or Director, chosen by the meeting, shall preside At all other meetings of the Institute the President, or, in his absence, one of the Vice-presidents shall preside If none of these be present the meeting shall elect a Chairman The President shall do and perform such other duties as may from time to time be assigned to him by resolution of the Board of Directors of the Institute II ORDER OF BUSINESS At each business meeting of the Institute the order of business shall be as follows 1 Reading of minutes of preceding meeting 2 Report of the President 3 Report of the Treasurer 4 Report of the Secretary 5 Election of Directors 6 Reports of Standing Committees 7 Reports of Special Committees 8 Special Orders 9 Miscellaneous business This order of business may be changed at any meeting, by a vote of a majority of the Members and Associates present At all sessions of the Institute, other than business meetings, the order of proceedings and the time of adjournment shall rest in the discretion of the presiding officer

Jan 1, 1923

By Alexander Gow

THE object of this paper is to discuss the fundamental principles of ball milling and to present some observations which have been made in laboratory and plant investigations. The discussion will be limited to wet grinding of ores in cylindrical mills, but doubtless many of the principles involved will be equally applicable in the other fields of fine grinding in which mills of the same type are used. In the concentrators of the United States there are about 650 ball, rod, or pebble mills, operating or ready to operate, and these have about 100,000 connected horsepower. For grinding in cement plants there is about three times this connected horsepower, and other industries have an unestimated amount, all used for the general type of grinding that is to be discussed here. It is true that these figures do not represent stupendous amounts of power, but when one considers that ball milling is often the "bottle neck" in a concentrator and that it often accounts for the largest single item in the cost sheet, the importance of a compre-hensive understanding of the principles of ball milling can be appreciated. Much of the past literature devoted to the subject has been concerned with the Kick-Rittinger dispute over evaluating grinding, or with dissertations on ball paths and the closely allied impact versus attrition hypotheses, or with the exclusive consideration of certain specific varia-bles, such as mill speed or circulating load, irrespective of the other factors in operation. Such subjects may be dismissed with but little discussion in what may be called, by comparison, a new philosophy in the study of ball milling. Briefly, this "new philosophy" consists of using power as an index of the mill operation, and investigating grind-ing from the standpoint of set and induced variables, with emphasis on the latter.

Jan 1, 1934

By Oliver Hopkins

As prefatory to the body of this paper, a few general statements will be made (1) in regard to the history and importance of the asbestos industry, (2) as to the principal sources of the raw material, and (3) as to the types and modes of occurrence of asbestos. While the modern asbestos industry is comparatively new and its growth during the last decade or two has been marvelous, asbestos has been known and used in a small way for probably 2,000 years-at least since the time of the Roman Empire. Pliny refers to it as "the funeral dress of kings" and Plutarch records its use as lamp wicks. In the middle of the 18th century the deposits of the Ural mountains were opened up and the first factory for the manufacture of asbestos goods was established. But, owing to the limited demand for the goods, the industry disappeared and interest lapsed until about 50 years ago. Since 1860, the asbestos industry has made wonderful strides, and each year sees new uses to which the material can be put, and hence an increased demand and an increased output. Of the three great producers of asbestos, Canada, Russia, and the United States, Canada stands pre-eminent, producing from one-half to three-fourths of the world's supply; Russia comes second, producing ,in 1911 about two and one-half times ,is much as the United States. In 1912 the Canadian production,1 including both asbestos and asbestic, was 131,260 short tons, valued at $2,979,384; while during the same year the production of the United States2 was 4,403 tons, valued at $57,959. Canada holds the first place in the production of raw asbestos, while the United States holds the first place in its manufacture.

Jan 9, 1914

By John D. Price

COAL blending, in the preparation of coal before coke making, is so commonly practiced as to be almost universal. But the reasons underlying this practice, the benefits resulting from it, and the materials used in blending vary widely. This paper will outline the various phases of the subject and present information which may be correlated with work that has been done elsewhere. It will deal entirely with work done on the high-volatile coking coals of the western part of the United States, special emphasis being given to the coals of Colorado and Utah. A surveyL of the 86 coke plants in active operation in the United States during 1949 indicates that only 9 plants, or 10.5 pct of the total, charged one single rank of coal into their ovens, while the remaining 89.5 pcl made use of blending in some form. This report indicated that of these total plants 5 used straight high-volatile coal, 4 used straight medium-volatile coal, 47 used blends of high and low-vola-tile coals, 25 used blends of high, medium, and low-volatile coals, 2 used blends of high and medium-volatile coals, 3 used blends of medium and low-volatile coals. The fact that certain plants operated on a single kind of coal should not be interpreted to mean that no blending was practiced there, for invariably such plants secure their coal from more than one source and in the interest of uniformity do blend the coals as received. The general term coal blending covers two fields, the first of which is the mechanical mixing of a number of coals to secure uniformity. Often it is found necessary to secure coal for coke production from a number of different mines; these coals, though of the same general type or rank, may differ in their chemical composition or in the physical qualities they impart to coke made from them. Again, it is not unusual to find that coal from different sections of the same mine may show variations in quality. Under such conditions it may be necessary, in the interest of a uniform final product, to introduce a system of blending bins, a bedding yard, or other mechanical methods of securing a uniform mixture. Unfortunately this form of blending has received very little attention up to the present time; it has not received the consideration its value merits. The second type of blending, while also for the purpose of coke improvement, deals more particularly with the use of a blending agent differing in character from the base coal: it is this form of blending that will be discussed here. To consider only the western coals, for blending may be found necessary for other reasons with other coals, blending has been practiced experimentally or commercially under the following conditions: 1—When a single coal or mixture of coals of the same rank does not produce a satisfactory coke. For example, a high-volatile coal when used alone is likely to contract when coked so that a comparatively weak coke is formed. Or, if of very low rank, the coal may be deficient in the necessary bitumens required for good coke production. 2—When a product of some special quality is required, for example, when a plant ordinarily producing blast furnace coke must operate at slow coking rate to produce a high-grade foundry coke. Under this condition the reduced daily production of all products which accompanies slow coking time may be undesirable, and the use of some blending agent to increase the size of coke made at faster coking rates may be necessary. 3—When greater yield of coke or its coproducts is needed. Depending upon economic values of the products it may be found desirable to increase the yield of one or the other. 4—When supply of a particular coal must be used, either to protect reserves of high quality coking coal or to utilize a surplus or inferior product not otherwise usable. Many materials have been used for blending purposes, the exact agent to be used depending both upon the condition to be corrected and the nature of the base coal. No universal blending agent that can

Jan 1, 1954

By Benjamin B. Lawrence

Arthur Launcelot Collins was born at Truro in Cornmall, England, July 8, 1868. His father, J. H. Collins, is well kuown as an engineer, and his books have been widely read in this country: Arthur was the third son. Three of his brothers are engaged, as is his father, in the profession of mining engineering. His education, gained partly at St. Austell, partly at Falmouth, and partly at Nuttall's Collegiate School in Truro, England, was supervised by his father, who was then Lecturer on the Mining Sciences to the Miners' Association of Cornwall and Devon. In 1881, J. H. Collins went to Rio Tinto, in Spain, as chief chemist and metallurgist, and it was there that Arthur began his practical work in the laboratory and continued his studies in chenlistry and mineralogy. To the latter science he was deeply attached; and he never lost an opportunity of adding to his cabinet of minerals, of which he was an ardent collector. In 1883, he became assayer and general assistant to Mr. Nicol, Manager of the Peña del Hierro copper-mines, ncar Rio Tinto. Until 1885 he was associated with his father in the study of the geology of the Rio Tinto belt. In 1885 he retuined

Jan 1, 1904

By D. G. Brubaker

IN the history of geophysical exploration by the electromagnetic method many procedures and types of equipment have been used. Source arrangements for surveying on the ground have included long wires grounded at both ends, large loops laid on the ground, small horizontal loops, and vertical loops of moderate to small sizes. Receivers have ranged from simple direction-finding coils to combinations of coils and a bridge circuit to measure quantities related to both amplitude and phase of the received signal. Primary considerations in choosing a survey method are effectiveness, speed, and economy. In work on the ground in which locating and out- lining conducting bodies is the objective, methods using a vertical coil and a simple direction-finder have been found the most satisfactory. Mason1 and his associates applied such a method by setting up a vertical source loop, fed with low frequency alternating current by an engine-generator set. With a direction-finding receiver they surveyed lines along the ground in the plane of the loop, taking measurements at each station of the azimuth and inclination of the magnetic field direction. Deviations of this direction from the normal direction, that is, from horizontal and perpendicular to the traverse line, indicated a conducting body. Recent advances in design of light-weight generating equipment have encouraged using this and similar methods.

Jan 7, 1957

By C. V. Millikan

When a flowing oil well is being drilled in with cable tools, it is difficult to determine when an additional streak of pay sand is drilled. By placing a gas meter on the gas-release line from the oil and gas separator, any increase in volume of gas that comes with oil production is recorded. In addition to the better knowledge of sand conditions, these data are a valuable aid in determining the size and depth of shots, plugging off water, and estimating ultimate production. The meter also gives a record of operation while the well is drilling in; and if allowed to remain on the well after completion will give information that may lead to a more efficient recovery of the oil from the sand. A VERTICAL cross section of any oil sand will show great variation in porosity and oil content. In some sands, the entire thickness contains oil, the variation being only the degree of porosity or oil content, but more often the pay streaks are distinctly separated by shells or barren streaks. These pay strata vary in number and may be thin or thick. The total thickness of the sand may be several tens of feet while the part containing oil may be only 1 or 2 ft. thick. The amount of production is not proportional to the thickness of the sand but rather the thickness of the pay streaks. If the total thickness of the sand or the amount that can be drilled is small, or if the initial production is small, the depth at which each stratum of pay is encountered can be determined sufficiently accurately for all present economic purposes. However, when drilling through sand of considerable thickness, giving large initial production, especially where conditions necessitate that the production go through the same gas trap with that from one or more wells already producing, it is difficult to determine the depth at which the pay strata are encountered even accurately enough to guide the placing of shots, and small pay streaks are apt not to be recognized at all.

Jan 9, 1925

By Godfrey T. Vivian

Very often important discoveries are made in one industry that may be used to advantage in another, but, owing to the rarity that men step out of one industry into another, these discoveries remain unknown outside of the trade or profession in which they are made. The present paper is written solely with the object of bringing to the notice of mining and metallurgical engineers a centrifugal ore-concentrator which eventually may have a great application. Ore- Grading. Fig. 1 represents the action which takes place when solids of different dimensions are stirred up in a box and allowed to settle as minerals of the same size but of different specific gravity. This principle is applied practically in grading china-clay and in enriching tin-slimes. In the latter case the slimes are stirred around in a tub with water and allowed to settle while at the same time the side of the tub is struck with a hammer. The colloidal slimes are then poured off and various slices taken off horizontally. The great drawback of this process is that the particles are deposited horizontally and all the colloids are lost. Fig. 2 represents the action in which the particles are deposited side by side. Here the drawback is the great length of the box needed in order to prevent excessive loss. The alternative to lengthening the box is to increase the force of gravity. Centrifugal force may be regarded practically as the force of gravity augmented. Fig. 3 illustrates the Gee centrifugal machine for grading ores. The drum, A, fitted with a box, B, is mounted on a shaft or spindle, C, the whole being suspended from a ball-bearing of special design at D, supported between girders, E. Rota-

Jan 1, 1913

By Bertrand S. Summers

The technologist who has devoted much attention to the foundry-business will perhaps have noticed that the present scientific development of foundry-practice is in a condition similar to that of the steel-business some twenty years ago. In many cases chemistry is looked upon with suspicion, and the old cutand-try-methods are still in vogue, especially in the smaller foundries. The larger works have laboratories and more extensive means for applying science, and have devoted considerable time and attention to the improvement of foundry-practice. This is more especially true in the East, although in the West there are numerous firms devoting considerable attention to improving their metllods through the application of chemistry to work in the iron-foundry. It is hardly necessary to dwell extensively upon the effects of the different metalloids associated with iron, in pig- or castiron, in the way they are given in most text-books. The writer has thought proper to consider more particularly new developments in eupola-practice, particularly those pertinent to the discussion of the physics of cast-iron. In this connection some new ideas which recent practice has developed will be mentioned, and some results of practical tests along the same line will be described. Yet it has been deemed advisable also, by way of introduction, to outline the effects of the metalloids as they are described in most text-books, without going into detail. The question is sometimes asked, What is the most important element governing the quality of cast-iron 'for any particular purpose ? It would seem that right here lies the basis for the greatest difference of opinion. Some twenty years or more ago, silicon was regarded by foundrymen as one of their worst

Jan 1, 1899

By Charles H. Snow

In vol. xxxviii. (1889) of the American Journal of Science, under the heading " Pseudomorphs of Native Copper after Azurite from Grant County, New Mexico," Mr. W. S. Yeates describes a most interesting mineral form. His account is, that a material having the appearance of native copper, but brittle and much lighter than native copper (its specific gravity being but 4.15), was submitted to him by the United States National Museum for examination. The substance was reported to him as occurring in masses, varying in weight from one ounce to seventy pounds, and composed of tabular crystals, variously grouped. The specimens, when submitted, ex hibited on the outride a covering of clay; and they were subse quently found to consist of particles of the same clay, intimately mixed with others of native copper, producing the aspect, without the weight, of pure copper. The forms of the azurite crystals, found at Bisbee, Arizona, having been suggested by the specimens, actual goniometer-measurements were made; and the two crystallizations were thus proved to be identical. To these statements I may be permitted to add some further in formation, derived from personal examination, and illustrated with a suite of characteristic specimens collected by myself (Fig. 2).* Grant County, New Mexico, in which these specimens were found, is a region noted for variety of mineral species. The sixth edition of Dana's Mineralogy mentions many of the silver, copper, and lead minerals, and also turquois, aragonite, wulfenite, barite, fluorite, vanadinite, etc. I have noticed also green and red garnet, molybdenite, native alum, smithsonite, bornite, etc., some of which presented peculiar modes of occurrence. The green garnet, for instance, was in masses formed of separate crystals, each of pea-size and not unlike that vegetable in tint and smooth opaqueness. The crystallization was

Jan 1, 1893

By W. J. Reagan

All of the material described in the following paper is within the following specifications: carbon, 0.50 to 0.85 per cent; phosphorus and sulfur, 0.04 per cent max.; manganese, 0.60 to 0.85 per cent; silicon, 0.20 to 0.30 per cent. Steels with much lower carbon contents can be handled in exactly the same manner. One of the first things necessary to basic open-hearth slag control is uniformity of charge, particularly as to the silicon content and the percentage of lime charged. Our past work has shown that the silicon content of the charge is more important than the carbon content in determining the carbon at which a heat will melt. We have found that by keeping the silicon content of the charge about 0.60 to 0.65 per cent, and with a lime charge of about 7 per cent (limestone equivalent), our heats melt sufficiently high in carbon to allow a moderate feed of ore and the slag will be of low iron oxide content. In making a heat of 0.75 carbon steel we endeavor to have the heat melt at about 1.20 per cent carbon. The iron oxide content of the finishing slag depends upon the final phosphorus desired; for a given ratio of lime-silica in the slag, the lower the phosphorus desired in the final steel, the higher the iron oxide content of the finishing slag must be. By reducing the lime charge to a minimum, the heats melt faster, and by keeping the silicon content at the desired figure the heats melt with a high SiO2 slag, which has been shown in previous work to give a low iron oxide slag. In most of our work, using a 180,000-lb. charge and with basic pig containing about 1.20 per cent silicon, we use a lime charge of 8500 lb. of limestone and 2000 lb. of burnt lime. The pig charge is about 48 per cent of cold pig, and the charge contains a large percentage of works scrap. This causes a heat to melt with about 1.20 to 1.30 per cent carbon, for a heat that will finish about 0.75 per cent carbon. This gives a heat that will need a feed of about 2000 lb. of ore, and also a slag with an FeO content of about 4 to 5 per cent at melting. The additional FeO obtained from the ore addition will bring the FeO in the slag at tapping to about 6 to 7 per cent. It is necessary to carry quite a fluid slag to prevent pickup of FeO, as a too heavy slag causes FeO to build up in it, giving a finishing slag with a greater FeO than is desired. By adjusting the lime charge of a basic open-hearth heat, a very sensitive adjustment of the slag may be obtained. The figures in Table 1

Jan 1, 1935

By W. J. Reagan

All of the material described in the following paper is within the following specifications: carbon, 0.50 to 0.85 per cent; phosphorus and sulfur, 0.04 per cent max.; manganese, 0.60 to 0.85 per cent; silicon, 0.20 to 0.30 per cent. Steels with much lower carbon contents can be handled in exactly the same manner. One of the first things necessary to basic open-hearth slag control is uniformity of charge, particularly as to the silicon content and the percentage of lime charged. Our past work has shown that the silicon content of the charge is more important than the carbon content in determining the carbon at which a heat will melt. We have found that by keeping the silicon content of the charge about 0.60 to 0.65 per cent, and with a lime charge of about 7 per cent (limestone equivalent), our heats melt sufficiently high in carbon to allow a moderate feed of ore and the slag will be of low iron oxide content. In making a heat of 0.75 carbon steel we endeavor to have the heat melt at about 1.20 per cent carbon. The iron oxide content of the finishing slag depends upon the final phosphorus desired; for a given ratio of lime-silica in the slag, the lower the phosphorus desired in the final steel, the higher the iron oxide content of the finishing slag must be. By reducing the lime charge to a minimum, the heats melt faster, and by keeping the silicon content at the desired figure the heats melt with a high SiO2 slag, which has been shown in previous work to give a low iron oxide slag. In most of our work, using a 180,000-lb. charge and with basic pig containing about 1.20 per cent silicon, we use a lime charge of 8500 lb. of limestone and 2000 lb. of burnt lime. The pig charge is about 48 per cent of cold pig, and the charge contains a large percentage of works scrap. This causes a heat to melt with about 1.20 to 1.30 per cent carbon, for a heat that will finish about 0.75 per cent carbon. This gives a heat that will need a feed of about 2000 lb. of ore, and also a slag with an FeO content of about 4 to 5 per cent at melting. The additional FeO obtained from the ore addition will bring the FeO in the slag at tapping to about 6 to 7 per cent. It is necessary to carry quite a fluid slag to prevent pickup of FeO, as a too heavy slag causes FeO to build up in it, giving a finishing slag with a greater FeO than is desired. By adjusting the lime charge of a basic open-hearth heat, a very sensitive adjustment of the slag may be obtained. The figures in Table 1

Jan 1, 1935

By Herbert Haas

The following notes describe a laboratory for metallurgical chemistry and technical analysis which I built late in 1903, while engaged, as constructing engineer, in erecting a pyrite smelter at the Afterthought mine, Ingot, Shasta county, California. This design partly supplements the paper of Mr. C. H. White, which was presented at the Atlantic City Meeting of the Institute in February, 1904.' The work of a chemist at a busy smelting-plant differs radically from the work of a student, who takes all possible care to have his analysis exact within hundredths of a per cent., and has plenty of time to devote to the operation. The works-chemist, having the shortest time in which to make his determinations, should have a laboratory arranged as conveniently as possible to save time in carrying out his work. By a conveniently-arranged laboratory, I do not mean an elaborately furnished one; but rooms suitably fitted up and supplied with the necessary apparatus to do the work expected from the chemist within the requirements of technical analyses. It is not uncommon at large smelters to have the night ae-sayer prepare 30 or more samples for titration, in time for the

Jan 1, 1905

By Donald C. Moore

INTRODUCTION Published laboratory procedures used by the major crusher manufacturers for estimating crusher size are based upon techniques developed in the mid 1940's. Specifically, the two largest U.S. crusher manufacturers are using the Bond impact procedure. Although the procedure is still in use, there is little confidence in the results. Unless the laboratory results fall within a preconceived range, personal experience is used for selecting the number of crushers and connected horsepower (Coleman, 1978). There have been numerous theoretical studies of the communition of particles. However, no procedure has been widely acclaimed for predicting crusher power requirements. Experimental studies have included breakage by slow compression, slow speed impact and high speed impact. Theoretical results have increased our knowledge of the comminution process, but the procedures have not been capable of analyzing the complex mechanism of industrial crushing machines. The purpose of this chapter is to review the empirical crushing procedures that have been used, including advantages and disadvantages. A new procedure which has been in development at Allis-Chalmers Corp. from 1977 to 1982 is detailed. Actual results of the new procedure on a concentrator feasibility study are included. EXPERIMENTAL GOALS The goals of a laboratory crushability procedure are accurate predictions of: (1) Rock crushability, measured as energy per unit mass. (2) Crusher product size analysis. A procedure capable of predicting these parameters should be inexpensive and require a minimum amount of sample. CRUSHABILITY TESTS Bond Impact Procedure F. C. Bond (1946) detailed the development of the pendulum impact device that is still used by the leading crusher manufacturers (Bond; Rexnord, 1976). The device (Fig. 1) is comprised of two bicycle wheels with 13.6 kg steel hammers tangentially attached to the rims. The hammer faces are 50 mm apart at rest. On opposite ends of the hammers lifting

Jan 1, 1982

By T. F. Witherbee

ON the 22d of November, „1879, white iron unexpectedly appeared while working the Cedar Point Furnace, Port Henry, N Y., on the following burden, calculated to turn out mill and foundry iron: Anthracite coal, - 3000 pounds. Connellsville coke, - 1500 pounds. Total, - 4500 pounds. Old Bed ore, - 3750 pounds. New Bed fine ore, - 2000 pounds. New Bed chunk ore, - 1000 pounds. Kearnev hematite ore, - 750 pounds. Total, - 7500 pounds. Magnesian limestone, - 1865 pounds. Common limestone, - 885 pounds. Total, - 2500 pounds. The production per week since making mill iron-from September 1st, 1879, to date, seven weeks-was 412 gross tons. Ore was not immediately taken off, as nothing serious was looked for. The iron for the next week graded from No. 4 to white, and in. six days the make was 274 1/2 tons. At 10 P.M., Friday, the 28th, a slip of-..4 feet occurred, and on Saturday, at 5 A.M., another of 3 feet on the south side and. 7- feet on the north side took place. At this time it was found that the south side had been hanging for over a week--a fact which had not been reported. Throwing off the wind, to produce slips. and reduce pressure, had been tried with no effect. The stock, as seen through the tuyeres, seemed to consist almost entirely of coal, with just enough infusible cinder to cement it into a mass through which the blast could not be forced. At 3.50 A.M., Sunday, the 30th, the first cinder seen in seventeen hours came into tuyeres Nos. 1, 2, 3, and 6, At 9.30 A.M. No. 3 was taken out to see if the stock

Jan 1, 1881

By W. L. Maxson

GRINDING is one of the major problems in present-day milling practice, and in many cases, it represents one of the main items of expense. It becomes necessary from time to time, to compare grinding practice in one plant with that in another plant, and in making such comparisons, it would be very desirable were there available accepted standards of grinding resistance for the materials involved. Such a set of grinding constants would also be of considerable value in planning new grind-ing installations. There have been some tests published, notably those of Lennox,1 which have been generally cited as a basis for such comparisons. Most of the tests made have been based upon open-circuit grinding, in which a weighed sample of the ore was crushed to pass a certain mesh, and was .then ground under certain conditions, either wet or dry. Grinding was continued until a certain desired fineness was attained, and the compari-sons based on the grinding time, or the samples were all ground for a certain specified length of time and comparison based upon the screen analysis of the products. Such methods have been acceptable where the ore was to be ground in open circuit. However, in modern milling practice, many grinding problems involve closed circuits wherein the oversize returns to the mill and a large circulating, load builds up. A method of testing that does not take the circulating load into consideration might yield results quite different from those to he expected in practice, because the minerals in the ore that offer maximum resistance to grinding will accumulate in the circulating load up to an equilibrium value. Thus the nature of the ore charge in the mill at any instant may differ greatly from that of the new feed, with a consequent error in the rate of grinding determined by such a method.

Jan 1, 1933

By John J. Dran

The elaborate analytical techniques for risk analysis used by the large mining companies in conjunction with the valuation of mineral reserve sites are totally inappropriate for the small miner. The small miner does not possess the financial and technical resources which are available to the large companies for such risk analysis. Even if these resources were available to him, the relatively small size of his investment would not justify a huge expenditure for risk analysis. What is required by the small miner is an improvement over the single point estimates provided by the valuation techniques that he currently uses. Such improvement should provide an indication of the extent to which the true value (known only after the fact) might differ from his estimated value of the mineral reserve site. In addition, a rough indication of the chance of occurrence of various differences between the true and estimated values would be helpful to him. These improvements must be obtained at low cost to be useful to the small miner. Such improvements would have to be based on a simple system requiring very few additional estimates by the miner. The small miner is not accustomed to dealing with sophisticated risk analysis techniques. Hence, any complex, time-consuming approach would be doomed to disuse or possible misuse. A simple, uncomplicated approach requiring a minimum of additional estimates by the miner can be based on a system devised by Frederick S. Hillier Although Hillier's method can be applied to various mineral reserve valuation methods used by the small miner, this note is specifically oriented to the discounted cash flow model. The basic discounted cash flow (DCF) approach provides a single point estimate of value and introduces the concept of risk (i.e. the possibility that the annual cash flows may differ from their estimated values and thus lead to an incorrect investment value) through the use of a risk-adjusted discount rate. This approach simply lowers the present value estimate with increases in the discount rate (i.e. increases in the risk assessment). The single point estimate still remains. Although the lower value estimate makes it less likely that the risky investment will be deemed profitable and undertaken, it provides no indication of risk that is suitable for examination and analysis. In addition to the "best guess" estimate of annual net cash flows required in the DCF model, application of the Hillier risk analysis model requires that the miner

Jan 1, 1976