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Gentlemen.-It is pleasant to be able to report that the year just closed has been a prosperous and successful one for the Institute. As is shown by the reports of the Treasurer and the Finance Committee, the business affairs of this organization are in excellent condition. With property valued at more than three-quarters of a million dollars, and with considerable sums in various funds and endowments, it is possible to spend on each member of whatever class $22.65 each year, more than the maxi- mum amount of dues received from any one, and substantially twice the average amount, $10.94, received from all. To do this has required close management and a strict regard for our limitations. In 1926, when the Budget was made up, the estimated expenditures for the work which it was felt the Institute must do exceeded the estimated receipts by some $14,000. At the close of the year, despite a number of unexpected decreases in receipts, it had proved possible to accomplish substantially all the work laid out and close the books with the usual transfer of initiation fees to the General Reserve, and still only reduce the cash balance $2,074.54. This was made possible by the hearty and loyal cooperation of the whole staff of employees and the extremely large amount of gratuitous service rendered by the officers, the members of the Divisions and technical Committees, and the membership generally. Even so it was impossible, because of limitations of funds, to undertake many things it is highly desirable for the Institute to do, and it was necessary to refuse the publication of important papers and committee studies, because of the lack of funds. PUBLICATIONS In an analysis of the income and expenses of the Institute published in the February issue of our Journal, MINING AND METALLURGY, it was pointed out that the largest item of expenditure is, as it should -be, for publications. Printing, binding, and distributing the various technical papers sent to the members cost in 1926 more than all the money received from dues, including the portion allocated as subscription to the magazine. Within the. year, the following volumes were printed and, with the exception of the second volume of TRANSACTIONS, which was mailed in January, were distributed : TRANSACTIONS, Volume 73, 1241 pages, covering papers on non-ferrous metallurgy, milling and concentration, non-metallic minerals, iron and steel, and the Institute of Metals Division papers read at the

Jan 1, 1928

By E. S. Bardwell

Each of the three refining furnaces in use at Great Falls is provided with a waste-heat boiler. The general arrangement of furnace and boiler is as shown in Fig. 1. Two of the furnaces have hearths 45 ft. 7 in. by 14 ft. 6 in.; the third has a hearth 46 ft. 101/2 in. long and of the same width as the other two. The furnaces at present are equipped for firing with natural gas. As originally constructed two of the furnaces had hearths 36 by 14 ft., and were designed to use coal on grates, but these furnaces were lengthened and fired successively with pulverized coal and oil. The third furnace was added in 1926 and was oil-fired until 1928. Since that year natural gas has been in use on all of the furnaces. The changes that have occurred in connection with furnace design and the comparative performance of the furnaces with various fuels were described1 by the author in 1932. Two of the furnaces are provided with Stirling class A-30 boiler. rated at 400 hp. and having 4008 sq. ft. of heating surface. The thir furnace has a class A-21 Stirling waste-heat boiler rated at 300 hp. with 2805 sq. ft. of heating surface. The boilers are not baffled. The waste-heat boilers are placed on a platform 17 ft. 3 in. above the level of the furnace charge floor. The waste gases from the furnace pass by a vertical uptake flue, 4 by 3 ft. inside at the point where it enters the chamber, to a brick chamber from which the gases pass over a bridge wall and into the boiler setting. The uptake flue and chamber are constructed of silica brick. Each furnace has a steel stack 125 ft. high, measured from the boiler floor. The stacks have a 41/2-in. firebrick lining and are 5 ft. in diameter

Jan 1, 1934

By R. R. Wilson

THE Cassidy Colliery operated by the Granby Consolidated Mining, Smelting & Power Co., Ltd., is situated about 9 miles in a southerly direction from the city of Nanaimo on Vancouver Island. The coal seam worked, which is called the Douglas bed, outcrops in the western part of the property, clipping easteyrl, and is opened by a slope 7 by 14 ft. in the clear, and is timbered by 12 to 14-in. framed sets of timber, 4 ft. center to center. A separate manway provides for a traveling road and at the same time forms an intake airway. The general average of the dip of the seam at Cassidy is 18°, and varies in thickness from 1 to 25 ft., averaging about 8 ft. The main slopes follow straight down the dip a distance of 4545 ft. At the bottom of the main slope two rock tunnels were driven a distance of 1500 ft. to the south, cutting through a fault, and opening up an area in the same seam which is known as No. 7 South Section. The coal is found in this section at a vertical depth of 1560 ft., and is very soft and friable. The immediate roof is a strong, sandy shale overlain by massive sandstone while the floor is a light-colored shale, hard when freshly exposed. It does not soften materially oil exposure. In the folding of the coal measures a great thrust caused a movement along the coal seam as indicated by the slickensided nature of the coal. The floor has buckled and this has rolled and squeezed the coal-so that with rare exceptions it shows no normal bedding but curved slip planes, which frequently start at right angles to the floor, curving upward and then in the direction of the dip. As such planes are slickensided it is probable that there is a coating of clay which is more or less impervious to gas.

Jan 1, 1927

By Nelson Severinghaus

FOLLOWING the end of hostilities of World War II, a resurgence of industrial and home building has given impetus to the use of many products of the stone industry. Shortage and high prices of some building materials such as lumber have aided wider acceptance of materials such as concrete products which have been relatively plentiful and which have not risen so rapidly in price. Ingredients for many of these substituting materials come from the stone industries, often from by-products of stone operations. Most of the tonnage of the stone industry in the State of Georgia comes from granite and marble. There are extensive dimension stone operations in granite in the Elberton and in the Lithonia areas. Building and monumental marble from the Tate area has been used throughout the country. Cutting operations in both types of stone have their by-products in faulty blocks and in spalls. In the State there are three large crushed-stone plants utilizing granite, one near Camak, one near Lithonia and one at Stockbridge. All have had excess fines or screenings in the past but this excess is largely disappearing with development of new uses. At times this "waste" has amounted to as much as 20 pct of the stone quarried. Handling of this large tonnage from point of screening out to storage piles presented a costly problem and large areas were taken up by the piles. Crushed stone is generally sold with a relatively narrow margin of price over cost, so any elimination of disposal cost or increase of income by marketing of by-products is highly desirable. BY-PRODUCTS OF GRANITE DIMENSION STONE It sometimes becomes difficult to differentiate between main product and by-product. In the cutting of granite dimension stone in the Lithonia district of Georgia, a considerable quantity of remnants of uneven shape and size is produced. The larger of these pieces have long found use as building rubble, particularly for retaining walls and building foundations (Fig I). They are broken down to one man size by plug drilling and wedging and by hammer. In recent times, demand for this former by-product has grown so that quarries have been opened for the production of rubble as a main product. Local masons have attained skill in using these irregular shapes to form economical and attractive walls. Areas such as Elberton, Georgia, which do not have a large building market nearby, tend to accumulate piles of large pieces of reject stone. When a road job or other local construction offers a quantity market for crushed stone, a portable crusher works this material down to crushed aggregate. Spalls and finer chips from granite stone-cutting operations have generally found their way into crushers and thence to the many uses for crushed aggregate. One company near Lithonia started marketing the smaller by-product crushed sizes for chicken grit about ten years ago. This use has grown to nation-wide distribution and the tonnage is now so great that they have opened a large quarry with

Jan 1, 1947

PLANS for the World's Engineering Congress in Japan are rapidly taking shape. The Congress itself will be held in Tokyo in the week beginning Oct. 30, and will be followed by a second week of short excursions to points of interest, and this, in turn, by a two weeks' excursion through the Inland Sea and to Manchuria. Aside from the time each may desire to spend in Japan, from 30 to 42 days, according to route, must be allowed for travel from New York to Tokyo and return, with correspondingly' shorter time from intermediate points. The American Express Co., which has been appointed official travel agents by the American committee, has fur-nished the following estimates:

Jan 8, 1928

By Barnabus Bryan

THE normal statistics of supply and demand of petroleum and its products have little more than local influence on the market movements of oil securities. The two major reasons for this condition might be called the technique of production and the technique of refining. The speed of improvement in each and the complete lack of control of their, inter-relation have resulted in overproduction of both crude oil and gasoline in the face of immense in-crease in demand. The trouble with oil securities has been that of the mushroom radio companies, where too frequent changes have killed the possibility of profits. It is a case of too much efficiency. During the past ten years there has been continual talk of wasting reserves of crude oil. This has been based on "present methods of production." Yet the methods of ten years ago in the oil industry are as antiquated today as the kerosene lamp.. The fear of famine has transformed the industry. The refiner, through the cracking process, has doubled the gaso-line output from a barrel of crude oil, thus doubling reserves and changing a current undersupply into oversupply. When the automobile driver desires he can again double reserves and transform want into abundance by demanding an automobile which will deliver twice as many miles per gallon of gasoline. Entirely apart from refining and in complete dis-regard for its needs, the oil producer has forged ahead at a pace too rapid for his own good. The science of geology has developed with experience and' lavish expenditure into a constant danger to profits in the production of oil. Two and one-half per ;cent of the producing wells in the United States are yielding one-half of the oil, while their flush production is used as an excuse to force down the price of crude oil until 200,000 wells must be abandoned or produce at a loss. The fabulous yield of the new high-pres-sure fields enables the fortunate owners to make a profit regardless of the price of oil. Such a condi-tion is unhealthy from the standpoint of oil itself and as a national policy. It is manifestly absurd to suppose that such a condition can endure for a gen-eration. There will come a day when the sequence of discovery will be temporarily broken, and at that time the consumer will be forced to pay heavy interest on the foolhardy policy of today.

Jan 6, 1927

By Samuel G. Lasky

The writer has been interested in determining the mineralogic dis-t,ribut.ion of silver in the base-metal ore of thc Ground Hog mine of the Asarco Mining Co. in the Central mining district of New Mexico.' This ore consists of a varitextured mixture of sphalerite, chalcopyrite, galena and pyrite in a quartz gangue. The average ore contains 14 per cent of zinc, 9.5 per cent of lead, 5 per cent of copper, and 10 oz. of silver per ton. Generally it is accepted that the silver in ores of this type, in which no definite silver minerals can be recognized, is contained in the base-metal minerals. Published assays of pure mineral specimens from many districts show that each of the base-metal minerals contributes in some degree to the total silver content of the ores, and to determine the proportionate amount contributed by each mineral in any specific ore it has been the practice to assay selected specimens of each mineral, the purity of which has been determined microscopically. There being no such assays available for the Ground Hog ore, and it being impracticable to have enough of them made to give average results, the writer resorted to a mathematical attack upon composite assays of large quantities of ore and concentrates. Surprising results were obtained, and it is the threefold purpose of this paper to present these results, to describe the mathematical method used, and to point out the value of using this method even when assays of pure specimens are available. Value of the Mathematical Method The obvious way of attacking the problem would be to separate cleanly large quantities of ore into its component minerals and then to determine the average silver content of each mineral by assay of representative samples. Since an absolutely clean separation is not feasible, it has been the practice, as mentioned above, to assay pure specimens of the different minerals. No one will contend seriously that the silver content of a base-metal mineral is constant throughout any particular deposit, and it will

Jan 1, 1935

By Samuel G. Lasky

The writer has been interested in determining the mineralogic dis-t,ribut.ion of silver in the base-metal ore of thc Ground Hog mine of the Asarco Mining Co. in the Central mining district of New Mexico.' This ore consists of a varitextured mixture of sphalerite, chalcopyrite, galena and pyrite in a quartz gangue. The average ore contains 14 per cent of zinc, 9.5 per cent of lead, 5 per cent of copper, and 10 oz. of silver per ton. Generally it is accepted that the silver in ores of this type, in which no definite silver minerals can be recognized, is contained in the base-metal minerals. Published assays of pure mineral specimens from many districts show that each of the base-metal minerals contributes in some degree to the total silver content of the ores, and to determine the proportionate amount contributed by each mineral in any specific ore it has been the practice to assay selected specimens of each mineral, the purity of which has been determined microscopically. There being no such assays available for the Ground Hog ore, and it being impracticable to have enough of them made to give average results, the writer resorted to a mathematical attack upon composite assays of large quantities of ore and concentrates. Surprising results were obtained, and it is the threefold purpose of this paper to present these results, to describe the mathematical method used, and to point out the value of using this method even when assays of pure specimens are available. Value of the Mathematical Method The obvious way of attacking the problem would be to separate cleanly large quantities of ore into its component minerals and then to determine the average silver content of each mineral by assay of representative samples. Since an absolutely clean separation is not feasible, it has been the practice, as mentioned above, to assay pure specimens of the different minerals. No one will contend seriously that the silver content of a base-metal mineral is constant throughout any particular deposit, and it will

Jan 1, 1935

The writer has been interested in determining the mineralogic dis-t,ribut.ion of silver in the base-metal ore of thc Ground Hog mine of the Asarco Mining Co. in the Central mining district of New Mexico.' This ore consists of a varitextured mixture of sphalerite, chalcopyrite, galena and pyrite in a quartz gangue. The average ore contains 14 per cent of zinc, 9.5 per cent of lead, 5 per cent of copper, and 10 oz. of silver per ton. Generally it is accepted that the silver in ores of this type, in which no definite silver minerals can be recognized, is contained in the base-metal minerals. Published assays of pure mineral specimens from many districts show that each of the base-metal minerals contributes in some degree to the total silver content of the ores, and to determine the proportionate amount contributed by each mineral in any specific ore it has been the practice to assay selected specimens of each mineral, the purity of which has been determined microscopically. There being no such assays available for the Ground Hog ore, and it being impracticable to have enough of them made to give average results, the writer resorted to a mathematical attack upon composite assays of large quantities of ore and concentrates. Surprising results were obtained, and it is the threefold purpose of this paper to present these results, to describe the mathematical method used, and to point out the value of using this method even when assays of pure specimens are available. Value of the Mathematical Method The obvious way of attacking the problem would be to separate cleanly large quantities of ore into its component minerals and then to determine the average silver content of each mineral by assay of representative samples. Since an absolutely clean separation is not feasible, it has been the practice, as mentioned above, to assay pure specimens of the different minerals. No one will contend seriously that the silver content of a base-metal mineral is constant throughout any particular deposit, and it will

Jan 1, 1935

By Samuel G. Lasky

The writer has been interested in determining the mineralogic dis-t,ribut.ion of silver in the base-metal ore of thc Ground Hog mine of the Asarco Mining Co. in the Central mining district of New Mexico.' This ore consists of a varitextured mixture of sphalerite, chalcopyrite, galena and pyrite in a quartz gangue. The average ore contains 14 per cent of zinc, 9.5 per cent of lead, 5 per cent of copper, and 10 oz. of silver per ton. Generally it is accepted that the silver in ores of this type, in which no definite silver minerals can be recognized, is contained in the base-metal minerals. Published assays of pure mineral specimens from many districts show that each of the base-metal minerals contributes in some degree to the total silver content of the ores, and to determine the proportionate amount contributed by each mineral in any specific ore it has been the practice to assay selected specimens of each mineral, the purity of which has been determined microscopically. There being no such assays available for the Ground Hog ore, and it being impracticable to have enough of them made to give average results, the writer resorted to a mathematical attack upon composite assays of large quantities of ore and concentrates. Surprising results were obtained, and it is the threefold purpose of this paper to present these results, to describe the mathematical method used, and to point out the value of using this method even when assays of pure specimens are available. Value of the Mathematical Method The obvious way of attacking the problem would be to separate cleanly large quantities of ore into its component minerals and then to determine the average silver content of each mineral by assay of representative samples. Since an absolutely clean separation is not feasible, it has been the practice, as mentioned above, to assay pure specimens of the different minerals. No one will contend seriously that the silver content of a base-metal mineral is constant throughout any particular deposit, and it will

Jan 1, 1935

By Porter W. Shimer

The device here described has been found useful in sampling foundry-iron, and there is no reason why it should not be equally useful in sampling other metals, which are not too hard to be drilled with a breast-drill. It consists, as shown in Figs. 1 and 2, of an ordinary plumber's gas T (without bead) of about 1/2-inch in internal diameter, over the leg of which is snugly fitted a small tin cup, about If inches deep, to catch the drillings which are made by use of a breast-drill with a 1/4-inch twist bit, working through the arms of the T. The forward arm of the T is filed down to a blunt edge, over which is stretched a short piece of rubber tubing projecting a little beyond the end. The bottom of the inside of the front arm of the ? is filed down to a slope, to facilitate the passage of drillings into the tin cup. The rear arm of the T is fitted with a tin ring, projecting about 1/4-inch, for the purpose of holding in place a stiff spiral brass spring, which serves to press the sampler firmly against the metal to be drilled. The other end of the spring is supported against a shoulder of the breastdrill. The projecting rubber prevents the loss of any drillings, and also the falling of sand from above into the sampler. With ordinary care there need be no risk of getting particles of rubber into the sample. It is, of course, not possible to get a fair average sample of drillings, representing a pile of foundry-iron, by taking drillings from any one or two pieces, because of the frequently mixed character of the iron. This observation applies more particularly to the sampling of iron as it is broken and piled at the foundry, than to sampling at the works, where, if the furnace is working regularly, a few pigs may fairly represent a whole cast. By the use of this device, it is possible for a sampler to pass

Jan 1, 1901

By George B. Clark

More economical methods of rock disintegration are needed for mining and for rapid excavation. Two major systems are drill and blast and mechanical excavation, including tunnel-boring machines. Current drill and blast methods are cyclic and there- fore limited in speed of operation. Boring machines are limited to moderately hard rock due to bearing and cutter wear. The recent development of ANFO, which has reduced blasting costs markedly, was largely fortuitous because of the method of manufacturing AN prills. New slurries and fluid explosives are easy to load and have high energy-pressure characteristics.

Jan 1, 1971

By M. R. Goldick

THE Roan Antelope concentrator was originally designed with a nominal milling capacity of 6000 tons of copper ore per day but this was subsequently considerably exceeded. In broad outline the plant consisted of two No. 30 McCully gyratory crushers for coarse crushing, five 5 1/2-ft Standard Symons crushers in open circuit for secondary crushing, ten No. 98 Marcy ball mills arranged in five sections, each two-mill section with primary and secondary Dorr classifiers for two-stage grinding. One stage of flotation roughing was followed by two stages of cleaning, all in airlift type pneumatic flotation machines, with the final concentrates filtered on Oliver drum filters. Tailing was thickened in a 250-ft Dorr traction type thickener before pumping to the tailing dam. Concentrating operations commenced in June 1931, on the completion of the first two ball-milling sections and output was increased during the next few months as the remaining sections were finished. From the start, very few technical difficulties were experienced but a number of evolutionary changes have been made as a result of operating experience and the availability of improved equipment. These have resulted in plant and ore-flow alterations although the general scheme of ore treatment remains substantially as before. The main changes to date were the installation of a new crushing plant at Storke shaft, some 7000 ft west of the original Beatty shaft; the completion of a two-stage secondary crushing plant; an increase of 20 pct in ball-milling capacity; a change-over from pneumatic to a mechanical type of flotation machine for flotation roughing; the reduction of flotation cleaning from two stages to one stage. ORE TREATED Typically the ore treated consists of metamorphosed shales and sandstones with shales as the principal mineral-bearing rock carrying chalcocite, bornite and chalcopyrite as the main copper minerals. Chalcocite predominated in the original eastern end of the ore body but there is a pronounced trend to an increased proportion of the copper-iron sulphides as mining proceeds westward. Although the ore can be described as soft, this advantage is offset by the extremely fine dissemination of the copper minerals which calls for a high degree of comminution of the ore (90 pct minus-200 mesh Tyler standard) to give a reasonably satisfactory separation of mineral from gangue. Latterly, there has been a tendency for the ore to become harder as shown by a reduction in the ball-milling rate from a previous figure of over goo tons per ball-mill day to the present output of some 800 tons. Chalcopyrite is not so finely disseminated and has also proved more amenable to flotation than chalcocite. Nonsulphide minerals are mainly the so-called oxides such as malachite, cuprite and melaconite with some chrysocolla and

Jan 1, 1947

By P. S. Roller

P. S. ROLLER,*College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should he defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. x between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.," Cambridge, Mass. Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids., Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg' have ground

Jan 1, 1941

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead-producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741." Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832." Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead, the zinc, and the rock or waste materials. The first record of zinc production from the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1953

By C. O. Dale, W. J. Rundle

THE upper Mississippi Valley zinc-lead area was the first major lead-producing section in the United States. The lead ore, found near the surface in crevices, was relatively pure galena that could be smelted directly into lead, at first in log hearth furnaces and later in more efficient blast type furnaces. French Canadian fur traders encouraged the Indians to mine the lead ore and showed them how to smelt it into lead that had a high value for bullets1 Nicholas Perrot found lead ore on the Mississippi River bluffs near the junction of Wisconsin and Illinois and in 1690 established a trading post on the Wisconsin side of the river opposite the present site of Dubuque, Iowa.2 Shortly after 1720 discovery of Mine La Mott in Missouri diverted considerable attention from the Upper Mississippi area. Mining continued on a desultory basis with operations concentrated in the Galena, Illinois-Dubuque, area. In 1740 at least 20 miners were at work in the Fever River area around Galena and are reported to have shipped 2500 70-lb pigs of lead to Kaskaskia in 1741." Julien Dubuque established a mining and smelting operation in 1790 near the city that bears his name and was granted sole right to exploit the mining operations on the lands of the Sauk and the Fox Indians. He is reported to have produced 30,000 70-lb pigs of lead in 1805. Following the death of Dubuque in 1810 the Indians refused to let the white miners enter their lands, and little was done on the Iowa side of the river until the Indians were removed by treaty with the United States government in 1832." Early mining was entirely for lead but as the crevices were followed down, increasing percentages of zinc sulphide and zinc carbonate were encountered and at first discarded. Later a market became available for the zinc ores, and hand jigging devices were made to separate the lead, the zinc, and the rock or waste materials. The first record of zinc production from the area is for 1860. Production of zinc passed that of lead before 1900, reached a peak of 64,000 short tons in 1917, fell off rapidly and continually to about 2000 short tons in 1938, and since 1940 has ranged from 11,000 to 19,000 short tons. Lead has been of considerably less importance since 1900, and at present only about 10 pct as much lead as zinc is produced. Practically all of the zinc ore has come from orebodies that are rather flat and wide with considerable length as compared to width. Most of the early lead came from the crevice type deposit, but present production is from the predominately flat zinc orebodies. The Graham-Snyder orebody, scene of Eagle Picher operations, is practically all zinc with little or no lead being recovered. Marcasite, present in varying amounts, makes production of finished concentrates by gravity separation impractical. Satisfactory lead and zinc concentrates have been produced since flotation was introduced in the area in 1927. An acid recovery plant was operated for about 20 years after World War I, but it has been dismantled, and no recovery of the iron sulphides in the ores of the district is being made at the present time. In June 1950 there were three companies operating mines and mills, Tri-State Zinc Co., Calumet & Hecla Consolidated Copper Co., and Eagle Picher Mining and Smelting Co. The Vinegar Hill Zinc Co. had completed a shaft at a new orebody and had started to develop the mine which will supply the Cuba City mill. The Cuba Mining Co. was holding the Andrews Mine inactive. The Dodgeville Mining Co. was not operating but was exploring for additional reserves. Several small mines were selling ore to the Eagle Picher mill. A general area map is given in Fig. 1. The Eagle Picher Mining and Smelting Co. entered the area in 1946 with an active exploration campaign. Leases on a block basis were secured for the area south from the Wisconsin-Illinois line near

Jan 1, 1953

By J. D. Alexander, W. L. Martin, J. N. Dew

This paper presents data obtained using a flood-pot technique to determine the fuel available and the corresponding theoretical air requirements for in situ combustion of crude oils. Since the technique is relatively quick and easy, it is a practical and convenient tool for evaluating reservoirs as fireflood prospects. It is also a research tool which facilitates systematic study of the variables affecting fuel availability and corresponding air requirements. The understanding of these variables is of prime importance to those concerned with the technical and economic development of in situ combustion as an oil-recovery process. The experimental results show conclusively that the fuel available for in situ combustion is not a constant but, rather, varies with crude-oil characteristics, porous-medium type, oil saturation, air flux and time-temperature relationships. Thus, the fuel availability for specified field applications should be determined using actual reservoir crude and core material and the process conditions expected during in situ combustion in the reservoir. INTRODUCTION In situ combustion is a thermal process for recovering crude oil from reservoirs. The thermal energy released during the combustion of a small amount of the oil in place aids in the displacement of the remaining oil. Numerous articles have been published describing the in situ combustion process giving detailed results of laboratory and field experiments.10 In order to engineer an in situ combustion project, a number of important factors must be considered and determined. These factors include: (1) the amount of fuel consumed per unit of reservoir volume swept by the combustion zone, (2) the composition of the fuel consumed, (3) the amount of air required to consume this fuel, (4) the portion of the reservoir swept by the combustion zone, (5) the appropriate air-injection rates and pressures, (6) the amount of oil that will be recovered, (7) the rate of oil production and (8) the operating costs. Nelson and McNiell1 recently have described a procedure which utilizes laboratory combustion-tube data as a basis for the calculation of some of these design factors. Various authors have attempted to describe the in situ combustion process mathematically, and considerable progress has been made. Analytical solutions to the problem of heat transfer from a moving combustion front have been obtained for linear and radial systems."-' All of the published results involve the assumptions that: (1) fuel concentration is constant throughout the reservoir, or that fuel concentration is inversely proportional to the velocity of the front for a given rate of oxygen consumption; and (2) the fuel reacts instantaneously with injected oxygen, while liberating a constant amount of heat per unit weight of fuel at all temperatures. It seems both desirable and reasonable to test the validity of these assumptions experimentally. This paper presents laboratory data which were obtained by means of a "fire flood-pot" method for determining fuel availability and composition, and the corresponding theoretical air requirements for in situ combustion of crude oils under variable conditions. The mechanics of the method are similar to a conventional tube-run experiment.' The important differences involve the size of the reservoir model used and the method for providing the experimental environment. The new method subjects conventionally-sized core samples or unconsolidated sands to a programmed environmental sequence similar to that experienced by a similar volume of rock during the approach and passage of a combustion front in a long tube or in an oil reservoir undergoing in situ combustion. Restored-state samples can also be used. The small samples and relatively simple techniques involved allow an experiment to be set up, run and calculated in about three 8-hour days. This is a considerable improvement over long-combustion-tube techniques which can require several days to run and several more work days to set up and calculate. All the runs presented were run at 40-psig injection pressure. Pressure was not considered as a variable for these experiments, since we previously had found that it had only a small effect on fuel availability up to 600 psig.APPARATUS AND MATERIALS APPARATUS The fire flood-pot apparatus consists of a consolidated

By S. S. Solomons, V. H. Tillmann, L. Smaniotto

INTRODUCTION The Zinc Corporation, Limited and New Broken Hill Consolidated Limited (ZC/NBHC) mines are situated at Broken Hill, New South Wales, Australia (refer Fig. 1). Both companies are wholly owned subsidiaries of Australian Mining and Smelting Limited (AM&S). Annual production from both mines is 2.3 million tonnes of lead, silver and zinc ore mined by a variety of underground methods. From 1911 to 1968, conventional overhand cut and fill stoping (both open and square set timber) was used almost exclusively, though some trials with shrinkage and sublevel stoping were undertaken. The trial stopes used short holes in conparison with today's sublevel open stopes. Mechanised cut and fill methods with regular rockbolt support were introduced in the early 1960's and undercut and fill stoping replaced square set timber stoping for pillar extraction in the late 1960's. Since 1968, increasing emphasis has been placed on longhole open stoping (LHOS) although some areas of the orebodies are comrnrnitted to the abovemehtioned methods since existing stope development makes it impracticable to change. In 1979, approximately half the total production was mined by LHDS and it is expected that the LHOS proportion will increase in the future. A vertical crater retreat method of longhole mining has also been successfully introduced to replace the undercut and fill stoping of some pillars and it is expected that this stoping method will be used in some future primary ore extraction. This paper reviews the progress of LHOS technology at ZC/NBHC with emphasis on the stope planning and design process and the importance of rock mechanics applications. The objective has been to maximise the primary extraction of the orebody giving due consideration to stability and safety. This has led to the design of a novel sloped wall longhole stoping method which should eliminate the formation of pillars by mining the orebody in a "continuous" manner incorporating a sloped sand interface between the blocks of ore being mined and the previously mined sandfilled stope. The expected results include the complete extract of the orebody without the formation of pillars and with minimal dilution of ore with sandfill. The evolution of the method can be best illustrated by following the mining progress of the relatively new Southern 'B1 Lode orebody at the NBHC mine. GEOLOGY Southern '8' Lode is the southern extension of the 'Bt Lode orebody, which is currently being mined further to the north in ZC/NBHC (refer Fig. 2). The ore block being developed and stoped extends for a vertical height of 90 metres from No. 17 level down to approximately ten metres above No. 19 level. The ore minerals are sphalerite and galena with the zinc:lead ratio averaging approximately 3:l with a minor amount of silver associated with the galena. There is a sharp contact between the ore and the country rock on the footwall and keel, while the hanging wall gradates into medium and low grade material. The gangue is primarily quartz together with garnet, biotite, feldspar, gahnite, sericite and minor apatite. In various horizons within the orebody, grey feldspar, quartz and garnet pegmatites occur as stratiform layers. The ore and country rock are competent with some inconpetent boundaries.

Jan 1, 1981

By G. M. Schwartz, G. A. Thiel

THE first record of the quarrying of dimension stone in Minnesota dates back to 1820 when limestone was quarried locally for part of old Fort Snelling. Limestone quarries were operated at Stillwater, Mankato, and Winona as early as 1854. Granite was quarried first at St. Cloud in 1868, and within a few years thousands of tons were shipped to widespread points. Rough dimension stone for large buildings furnished the first important market, but beginning in 1886 paving blocks were in demand. The largest shipment was in 1888, when 1925 cars were shipped from the St. Cloud area. Quartzite was quarried first at New Ulm in 1859 and somewhat later at Pipestone and elsewhere in southwestern Minnesota: The productive dolomite quarries at Kasota were opened first in 1868 and have continued as large producers of a variety of stone to the present time. At present, the industry is controlled by relatively few operators, and for that reason detailed figures on dimension stone are not released for publication. A general idea may be obtained from the data in the Minerals Yearbook for 1948. The figures for total stone produced in Minnesota are 1,804,000 tons valued at $5,090,652. Probably the largest item in the latter figure is received from dimension stone. A better idea of the situation, in relation to the country as a whole may be gained by using the data for 1930 when more companies were operating in Minnesota, and complete figures were published. In that year Minnesota produced granite valued at $2,668,119 and ranked third among the states in value. Minnesota's production of granite was almost exclusively for dimension stone. In the same year Minnesota produced 300,000 tons of limestone (dolomite) valued at $840,860, and this likewise was mainly dimension stone. In finished limestone Minnesota ranked second among the states in 1930. Sandstone and minor amounts of quartzite are the only other dimension stones that have been produced in Minnesota, but the quarries are now inactive. The commercial stones of Minnesota have been described in two reports by Bowled and by Thiel and Dutton.2 The early history of quarrying in Minnesota and extensive notes on the various rocks are given by N. H. Winchell.3 Small limestone and dolomite quarries were numerous throughout the area of Paleozoic rocks in southeastern Minnesota. Early production was largely dimension stone. With the increased use of Portland cement, most of these ceased production, and today only those at Kasota and Winona remain in operation. In recent years many quarries have reopened and new ones started, but these are devoted to the production of crushed rock and agricultural lime. As the application of modern quarrying and finishing methods increased, small companies in the granite business have dropped out, and the remaining companies have modernized their plants, purchased old quarries, and opened up new ones, thus furnishing a wide variety of granites suitable for most of the customary uses. It is the purpose of this review to present notes on' the geology and operations of each of the quarries now operating within the state. Granites and Related Igneous Rocks The term granite as used in this report includes granites, gneisses, diorites, gabbros, and other igneous rocks. The granites of greatest economic importance are found in three widely separated regions, see Fig. 1. 1-Central Minnesota in the region of the city of St. Cloud, 2-the upper Minnesota River valley region, 3-the northeastern portion of the state, commonly referred to as the Arrowhead region. The St. Cloud Region: The rocks of the St. Cloud region are mainly granites and related rock types such as monzonites and quartz diorites. The stones may be grouped into three major types, namely, pink granite, red granite and gray granite. Most of the pink granite occurs in the area to the southwest of St. Cloud. The rock is best described as stone with large pink crystals set in a finer grained black and white background. The minerals of the matrix occur in remarkably uniform sizes, and the pink crystals are sufficiently uniform in. their dis-

Jan 1, 1952

By C. G. Dunn

WITH respect to theories of recrystallization in metals plastically deformed, it has been said that the present status of this subject is far from satisfactory.1 It may also be said that before any mechanism of recrystallization for metals and alloys can be set forth in a satisfactory manner, there must be made available considerable data on orientation changes that occur during the deformation and recrystallization processes. Although extensive data have been published on recrystallization in single crystals and in polycrystalline specimens, more results of a type that can be classified according to simple orientation changes would be desirable in this complex field. It might even be said that before classifications are made according to types of orientation changes, it may be possible to classify a considerable amount of data into certain kinds of orientation states if groups of common or nearly common orientation occur. Methods for computing the probability of occurrence for various types of groups have been published2 and it has been pointed out in particular that it should be possible to analyze orientation data for high-order twin relationships between old and new grains even when agreement is not at all exact. In this connection, work on recrystallization in silicon ferrite reported in a paper concerned with complex groups of annealing twins' indicated that some data were consistent with primary nucleation occurring according to a second-order or a third-order twin transformation. However, further work in this latter direction was definitely needed. It is, therefore, the purpose of the present paper to report data and analyses for primary nucleation in silicon ferrite following small deformations that have been carried out at the Pittsfield Works' Laboratory of the General Electric Co. Analyses have been made with the aid of standard groups of twins that are reported elsewhere .4 PREPARATION OF SPECIMENS AND EXPERIMENTAL PROCEDURE The material used in the present investigation was taken' from the same heat of steel as that used in a prior report.3 The finished strip in both cases was made by the Allegheny-Ludlum Steel Corporation according to methods that produce a high degree of preferred crystal orientation. Samples with large grains had the following composition: 3.45 per cent Si, 0.003 C, 0.008 P, 0.020 S, 0.075 Mn, 0.031 Al and 0.072 Cu. Thickness of specimens was 0.012 inches. Plastic deformation was produced by cold-rolling the samples on a mill with

Jan 1, 1945