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By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Masayuki Otagawa

THE mining methods adopted in Japanese mines are less known to the mining world than those of other countries, owing to the geographical remoteness, but they present many features of interest to mining engineers. The Ashio copper mine is at the village of Ashio, in Tochigiken, about 109 mi. (175 km.) from Tokyo by the Imperial Government railroad. The area of the concession owned by the Furukawa Mining Co. is 5000 acres (5023 ha.). The mine was formerly divided into three, viz.. Honzan, Kotaki, and Tsudo, which are distant 3 mi. (4.8 km.) from each other as shown in Fig. 1. However, in 1914, when the writer was general manager, the three mining plants were amalgamated into one with the result of simplifying the working projects and increasing the operat-ing efficiency. The mine was visited-by the Institute party in November, 1911, when Mr. K. Inouye was the general manager. Shortly afterward, the writer was appointed manager, and, in 1915, he was succeeded by Dr. K. Asano. At the end of 1917, Mr. K. Kibe became general manager and, after two years, was succeeded by the present manager, Mr. I. Sugimoto-. For some time Dr. J. Kojima was acting general manager. It is needless to say that the gentlemen mentioned are all members of the Institute, as is also Baron T. Furukawa, president of the company. Since the discovery of the enormous masses, called "kajika" or "bullhead" deposits; which the writer has described elsewhere,' the mining methods have been varied to meet the different conditions presented by these deposits.

Jan 2, 1922

By E. C. Pechin

The attention of the members of the Institute of Mining Engineers is asked to a description of the minerals of Southwestern Pennsylvania, as representing the minerals of an enormous area, stretching continuotisly, with local variations, along the western slopes, and at the base of the Alleghany Mountains, through Western Pennsylvania, West Virginia, and Kentucky, and destined to play no mean part in the future of the iron trade of the country. The particular locality is at and near Donbar, in Fayette County, where the most careful examinations and extensive developments have been made. The deep gorges cut by the Youghiogheny and Monongahela Rivers, in their passage through the mountains and lower lands, give admirable facilities for observing the measures lying between the sandstones of No. 11 and the upper coal measures. Apart from its varied mineral and agricultural resources this section of country has great historic interest attaching to it. It was down the Monongahela that Braddock led his ill-fated army to its defeat under the wails of Port Duquesne, and it was up the Youghiogheny River and Dunbar Creek that the shattered remnants hastily fled to gain the camp of Colonel Dunbar, the second in command, who was entrenched upon a high knob, at the head of the creek which bears his name, and which is still known and visited as Dunbar's Camp. Numerous streams, runs, and localities preserve the names of the guides and pioneers who accompanied, in early colonial times, young George Washington, who was pushing his surveys into the trackless and hostile wilderness, and many of them taking up large tracts of

By E. C. Pechin

THE attention of the members of the Institute of Mining Engineers is asked to a description of the minerals of Southwestern Pennsylvania, as representing the minerals of an enormous area, stretching continuously, with local variations, along the western slopes, and at the base of the Alleghany Mountains, through Western Pennsylvania, West Virginia, and Kentucky, and destined to play no mean part in the future of the iron trade of the country. The particular locality is at and near Dunbar, in Fayette County, where the most careful examinations and extensive developments have been made. The deep gorges cut by the Youghiogheny and Monongahela Rivers, in their passage through the mountains and lower lands, give admirable facilities for observing the measures lying between the sandstones of No. 11 and the upper coal measures. Apart from its varied mineral and agricultural resources this section of country has great historic interest attaching to it. It was down the Monongahela that Braddock led his ill-fated army to its defeat under the walls of Fort Duquesne, and it was up the Youghiogheny River and Dunbar Creek that the shattered remnants hastily fled to gain the camp of Colonel Dunbar, the second in command, who was entrenched upon a high knob, at the head of the creek which bears his name, and which is still known and visited as Dunbar's Camp. Numerous streams, runs, and localities preserve the names of the guides and pioneers who accompanied, in early colonial times, young George Washington, who was pushing his surveys into the trackless and hostile wilderness, and many of them taking up large tracts of

Jan 1, 1875

By Benjamin Gi ldersleeve

Uniquely colored, thin-bedded quartzite is quarried between Crossville and Crab Orchard in Cumberland County, Tenn. It is produced in all sizes up to the limits of transportation from beds usually ranging in thickness from less than 1/2 in. to 6 in. The stone is very resistant to wear and weathering and is marketed throughout the country for many different purposes in exterior and interior construction. The annual production exceeds 36,000 tons, valued at nearly $1,000,000.

Jan 1, 1950

By C. L. Berrien

Before the present company was formed, in 1916, each group of mines comprising the old organizations made its detailed daily and monthly mine cost records along the lines used before the consolidation; that is, there was no attempt made to standardize and compare the stope costs of all mines. The general cost sheets for all mines, however, were standardized, and cumulative labor cost sheets were compiled in the general superintendent's office. At most of the mines, a daily record or distribution sheet for each working place was kept, showing the number of men, hours of drilling, hours of timbering, hours of tramming, cars of ore, cars of waste, sticks of powder and feet of fuse. This information was kept by each shift boss and entered on the distribution sheet by him or, at some mines, by a tram boss. The superintendent solved his own problems under the direction of the general superintendent, but without coming in touch with other superintendents, either through records or general conferences. Probably 95 per cent. of the stopes were square-set worked by day's pay; there were some back-filling and shrinkage stopes. Many of the development places, such as shaft sinking, station cutting, drifting and raises had always been run by contract. These contracts were given, for no definite period, at so much per linear foot and were measured monthly, or when completed, the men being paid monthly. In some cases, fairness to the employee was not maintained. On the whole, there was no standardization. There were no stope contracts, because of the objections of the Western Federation of Miners to general contracting in the mines. Changes in Stoping Methods Early in 1916, W. B. Daly, then general superintendent, and the author, then assistant general superintendent, visited the Southwest mining districts with the object of finding some improved mining

Jan 1, 1923

By C. L. Berrien

Before the present company was formed, in 1916, each group of mines comprising the old organizations made its detailed daily and monthly mine cost records along the lines used before the consolidation; that is, there was no attempt made to standardize and compare the stope costs of all mines. The general cost sheets for all mines, however, were standardized, and cumulative labor cost sheets were compiled in the general superintendent's office. At most of the mines, a daily record or distribution sheet for each working place was kept, showing the number of men, hours of drilling, hours of timbering, hours of tramming, cars of ore, cars of waste, sticks of powder and feet of fuse. This information was kept by each shift boss and entered on the distribution sheet by him or, at some mines, by a tram boss. The superintendent solved his own problems under the direction of the general superintendent, but without coming in touch with other superintendents, either through records or general conferences. Probably 95 per cent. of the stopes were square-set worked by day's pay; there were some back-filling and shrinkage stopes. Many of the development places, such as shaft sinking, station cutting, drifting and raises had always been run by contract. These contracts were given, for no definite period, at so much per linear foot and were measured monthly, or when completed, the men being paid monthly. In some cases, fairness to the employee was not maintained. On the whole, there was no standardization. There were no stope contracts, because of the objections of the Western Federation of Miners to general contracting in the mines. Changes in Stoping Methods Early in 1916, W. B. Daly, then general superintendent, and the author, then assistant general superintendent, visited the Southwest mining districts with the object of finding some improved mining

Jan 1, 1923

By Robert G. Woolery, S. Ramachandran, James A. Weber, Donald J. Hansen

Union Carbide began looking seriously at heap leaching in 1971. At that time some 1.6 million tons of mineral averaging 0.40 kg U308 /t) (0.80 lb U308) were stockpiled at various sites around the Gas Hills, Wyoming, uranium facility. Several alternatives for economically exploiting these reserves were considered and heap leaching seemed the most attractive. At the time, Western Nuclear Inc. was operating a heap leaching project nearby in which the low-grade ore was leached with sulfuric acid, the product liquor treated in an on site solvent extraction circuit, and the stripped product liquor re-acidified and recycled. The operators were most generous in allowing us to visit their site and in answering our questions. They were also helpful in discussing the problems associated with heap leaching. Their experience confirmed our belief that it was possible to operate a heap leach profitably on material too low grade for conventional milling.

Jan 3, 1978

By L. J. M. Smits

A theoretical equation is developed describing membrane potentials of shaly sands as a function of the cation-exchange capacity per unit pore volume of the rock and saturating water salinity. This equation is based on the same simple physical model as that employed earlier7 describing the electrical conductivity of shaly sands. Experiments conducted with NaCl solutions showed that the equation is obeyed, except in the case of saturated salt solutions where minor deviations were observed. A possible explanation is given for this discrepancy. As the clay content and the porosity of shales may vary appreciably, shale membrane potentials also vary. It was found that membrane potentials of shales can be calculated accurately with the aid of the same theory. Consequently, changes in membrane potentials can be accounted for by simple measurement of cation-exchange capacities per unit pore volume of both shales and shaly sands. Charts are given to facilitate interpretation of the electrochemical part of the SP in terms of cation-exchange capacities and the salinities of mud filtrate and formation water. INTRODUCTION A spontaneous potential (SP) log is a record of the way in which the electrical potential of a sonde varies as it passes down a borehole. Such variations frequently reflect rock texture, e.g., permeable beds intercalated in an impermeable series often are indicated by a sharp change in potential. Nowadays it is recognized that SP logs may be interpreted in terms of formation-water salinity and formation shaliness. The four sources of the SP are illustrated in Fig. 1. El and E4 represent streaming potentials generated by the flow of mud filtrate through mud cake opposite permeable formations and through shale, respectively. An extensive review of the literature on the evaluation of these streaming potentials is given by Lynchl and will not be discussed here. The remaining part of the SP generally is called the electrochemical potential, which is divided into two parts: a membrane potential across the shale (E3, Fig. 1) and a membrane potential in the permeable sand at the diffusion boundary between formation water and mud filtrate (E2, Fig. 1). It was found by wyllie2 that the membrane potential across the shale, E3, is the most important contribution to the electrochemical potential, the shale acting as a semi permeable membrane. When the sand is clean it does not influence the diffusion process, so E2 equals the value of the so-called liquid-liquid junction potential observed when the two solutions diffuse freely into each other. When the sand contains clay, however, this interferes with the diffusion process and affects the diffusion potential generated. The greater the proportion of clay, the stronger this effect will be. Wyllie,3 Doll4 and Poupon et al.5 took this effect into account theoretically, starting from a physical

Jan 1, 1969

By W. M. Peirce

COMMENCING in 1922, each year a lecture has been presented to the Institute of Metals Division at this February meeting. The range of subjects has been very broad. Some speakers have dealt with the most advanced work in the very heart of the science of physical metallurgy. Others have stimulated our imaginations by discussing the most recent progress in related fields of science. Others have told us of the latest art, which is at any moment of greater practical importance than the latest science, of metals. I feel that this broad precedent in the choice of subjects will justify my choosing to talk about some of the prosaic details of how, during the last generation, industry has effectively organized research to accelerate progress in our own and all other fields of science and related art. A number of years ago I visited one of this country's well-known research laboratories and had the pleasure of chatting for a few minutes with its equally well-known director. Being particularly interested at the time in the organization problems connected with industrial research, I broached that subject. My host stated in effect that he didn't believe in organized research. He was, of course, resorting to a rhetorical hyperbole to point out the danger of overorganization which can so easily drift into regimentation and destroy individual initiative. Actually nearly all of us, much as we cherish our personal freedom of action, believe in organization because the history of progress in every field of human activity has been a history of men learning to work together, to cooperate, to create an efficient division of labor. We use organization to achieve this coordination. At the same time our form of organization must not stifle individual initiative or it will prove a deterrent rather than an aid to accomplishment. Research, too, must be organized for greatest achievement but especial skill and care are perhaps needed to avoid stifling the initiative of the individual workers. I have, of course, followed with keen interest and much benefit the vast amount that has been said and written on this subject in the past twenty years but I still find new stimulation in discussions with my good friends working in this field. It is my hope that this discussion may serve a like purpose. It would be ungracious not to make due acknowledgment to the many able contributors to this particular phase of human progress, but their very number and the great volume of their writing makes it impossible to trace to any one individual a particular significant and valuable idea. I shall content myself therefore with this general acknowledgment and with acknowledgment of the specific debt that every research man owes to those who have so ably kept before the public, and in particular before industrial executives, the place of research in the modern scheme of things. Because of the legitimate publicity thus given to research, the day is past when any

Jan 1, 1944

By L. J. Ingvalson, Roy H. Miller

IN 1948, as part of a program to expand the copper tube mill facilities of the American Brass Co. plant at Kenosha, Wisconsin, plans were formulated to convert the 100 ton capacity anode casting furnace at the Great Falls, Montana plant of the Anaconda Co. to the production of 3 in. phosphor -ized billets. The furnace was rebuilt completely and construction was started on an addition to the building. Members of the Great Falls staff visited various billet casting plants to draw upon the experience of other producers of this type of casting. Many features of other plants were incorporated into the Great Falls plant but, in addition, many new designs were made and perfected to make the plant the most advanced of its type. Construction was begun in 1949, with the intention of having the plant in production in the early part of 1951. In January 1951, as the plant was being rushed to completion, the government placed restrictions on the production of copper tubing. As a result, work was halted and the plant lay idle for two years. Interest was reactivated in May 1953 with the lifting of restrictions on copper. Plans were laid to go into production in July and the first charge was taken out July 1, 1953. Basically the plant is designed to produce from cathode copper a 3 in. diam billet, 50 in. long. weighing 110 lb, and containing 0.013 to 0.036 pct P, at a rate of 150,000 Ib per operating day. These billets are to be pierced and drawn into tubing. Furnace The furnace is a 100 ton capacity reverberatory type, gas fired furnace; 21 ft long and 9 ft wide. Gas is introduced to the furnace through two 38 hole Mettler gas burners. These burners are of chrome nickel alloy with four jets per hole. The furnace is charged with cathodes from the Electrolytic Copper Refinery plus return scrap from the casting operation by means of a Wellman Seaver Morgan Co. 7,000 lb capacity charging machine. The charge crane has a 20 ft boom and fork peel provided with an air ram to push material off the peel. The machine travels on overhead tracks and can be moved in any direction. The boom can be raised and lowered and swung through an arc of 180". With well stacked units of 6,000 lb, an experienced operator can put the first charge into the furnace in 40 min. After charging, the steel framed silica brick charge door is put in place and sealed with clay. After melting and skimming off the slag, the copper is blown until an oxygen content of 0.40 pct is reached. The metal is then heated to 2180°F, covered with coke, and poled. Poling is continued until a 3 in. diam test billet, 8 in. long, shows a flat set. The oxygen content is usually 0.030 pct at this time. At the end of poling, a temperature of 2160" to 2180°F is desired. In the side of the furnace opposite the charging door is the taphole slot, extending 36 in. upward from the bottom of the furnace. It is 4½ in. wide at the outside face and widens to 9 in. at the inside face of the furnace. It is filled with a wet mixture of sand, fireclay, and crushed coal. Flow of metal

Jan 1, 1957

By J. T. Lapsley, I. I. Cornet, A. E. Flanigan, R. Hultgren, J. E. Dorn

With the greatly expanding use of magnesium during the war, it appeared necessary to the War Metallurgy Committee that procedures of heat treating common magnesium casting alloys be investigated systematically. Accordingly, the National Defense Research Committee of the Office of Scientific Research and Development placed a research contract with the University of California, supervised by the War Metallurgy Committee. Research on the solution heat treatment and aging of magnesium alloy AZg2, (formerly ASTM No. 17; also known as Dowmetal C or AM26o) which was done under "Restricted" Project NRC-2I, in 1942-1943, and reported in detail in a final report to the OSRD on September 3, 1943, is the basis of this paper, which has been released for publication by the OSRD. American experience with magnesium alloys prior to World War II was limited, although some of these alloys had been in commercial use since 1909. Requirements of the aircraft industry for light alloys made it imperative to understand and utilize the advantageous strength-weight properties of magnesium castings. Magnesium casting alloys are commonly solution heat- treated to increase their ultimate tensile strength and ductility; this treatment may be followed by an age hardening at elevated temperatures, which raises the tensile yield strength but diminishes the ductility. In the present investigation, AZp was studied to determine the effects on properties of various solution heat treatment and aging schedules. The principles of heat treatment of magnesium alloys are the same as for other non-ferrous alloys. The alloy studied, AZp, has a nominal composition of 9 pct aluminum, 2 pct zinc, and minor constituents. Fig I shows, by the projection of isothermal sections on the concentration plane, the solubility surface1 of the ternary magnesium-rich solid solutions as determined by X ray 'analysis. From this figure it may be seen that above approximately 375°C (707OF), AZp alloy is a single phase alloy at equilibrium; below this temperature the equilibrium condition is a two phase alloy. The beta constituent which precipitates from the solid solution exerts a hardening effect. Because the zinc concentration is only 2 pct, ancl because of its solubility relations, AZqz resembles a binary alloy of magnesium with aluminum. The solubility curve1 of Fig 2 shows how rapidly the solubility of the beta phase decreases with decreasing temperatures. Unlike many aluminum alloys, magnesium alloys do not age harden at room temperatures; precipitation hardening of AZ² alloy is performed at about 177°C (350°F). At this temperature about 3 pct aluminum

Jan 1, 1949

By Herbert H. Kellogg

IN 1930, Taggart, Taylor and Knoll1found that addition of various electrolytes to suspensions of ground minerals resulted in the stopping or starting of Brownian movement of the suspended particles. On the basis of the many new facts found, they rejected the "classical "[t] theory of Brownian movement and proposed an "ionization" theory. This new theory was further supported in later publications by Taggart.3,4 The acceptance of the ionization theory by men in the field of flotation may be judged by the mention given this theory in Gaudin's "Principles of Mineral Dressing,"5 Wark's "Principles of Flotation,"6 as well as in a recent paper by Fitt, Thomas and Taggart.7 The proponents of the ionization theory have gone further than the mere statement of the new theory. A parallelism between the Brownian movement of mineral particles and their behavior in flotation was reported.1,3,4 Likewise, the correlation between flocculation-dispersion of mineral particles and the Brownian movement of these particles has been investigated.7 Bankoff15 has suggested that "slime-coating" can also be directly correlated with the Brownian movement of the slime particles. In each of the instances mentioned it is the ionization theory of Brownian movement that has been used to "explain" the correlations that were found. The essential conflict between the kinetic theory and the ionization theory can be understood best by a brief description of each. The kinetic theory* proposes that Brownian movement is caused by the impacts of liquid molecules against the solid particles. At any instant of time, the molecular impacts on all sides of the solid particle will not balance, and the particle will receive a push in some direction. According to this theory, all particles suspended in a liquid are in Brownian movement. However, if the particle is very large, the impacts on all sides will very nearly balance and the motion will not be noticeable. Another factor, besides the size of the particle, that affects the magnitude of the Brownian movement is the viscosity of the liquid-the greater the viscosity, the less the displacement of particles by Brownian movement. This theory has been verified experimentally by measurement of the velocity of Brownian movement, and by measurement of the compacting of suspensoid particles under the influence of gravity.2 The originators of the ionization theory found that the addition of various electrolytes to slime pulps caused cessation or commencement of Brownian movement, depending upon the chemical nature of the particle and the electrolyte. The Brownian movement of quartz particles was observed to stop upon the addition of certain electrolytes, where the conditions of the experiment prevented flocculation

Jan 1, 1945

By R. G. Knickerbocker

A STURDY and consistent expansion of the metal industry occurred in 1947 exemplified by an increase of approximately 30 per cent in steel consumption over 1946. For this major reason, ferroalloy metals have seen their most useful and important peace-time year. The current domestic rate of consumption of high-grade ferroalloy ores, together with added consumption of new foreign producers as well as for eign nationals' restrictions on high-grade exports, makes very evident the increasing difficulties of obtaining sufficient high-grade ores. This accentuates the need for new and less expensive methods of production of ferro-alloy metals of higher purity from lower-grade domestic resources. As an important factor in preparedness, procurement of these high-grade, -critical ferroalloy minerals and metals for our nation's stock piles is a most important task. Also, this stock-piling can be the means of supplying incentive for the industrial development of the new metallurgical processes and plants necessary for the recovery of ferroalloy metals from our domestic low-grade deposits.

Jan 1, 1948

By AIME AIME

CONCIDENT with what promises to be a record-breaking year of activity in the world's greatest iron-ore district is the announcement of plans to hold the regional meeting of the American Institute of Mining & Metallurgical Engineers at Duluth, Aug. 12 to 15 in¬clusive. More than 25 open-pit mines will be in operation in Minnesota's great iron-ore district this year, affording to those attending the convention an unusual op¬portunity of viewing practically side by side the interesting activities of power-shovel operations, railway haulage, truck haulage, conveyor-belt haulage, tower-excavator operations, iron-ore concentration, and underground mining. The ore can be followed from the mines to washing and concentrating plants and to the docks and lake carriers on Lake Superior, where an average of one boat of 10,000 tons' capacity will be loaded and depart every 15 minutes, day and night, during the 1941 shipping season.

Jan 1, 1941

By David R. Mitchell, E. A. Reilly, G. R. Gardner, F. P. Lasseter

THE methods that may be applied to the treatment of slurries and water, as these are related to practical coal-preparation problems, are concerned essentially with the movements of solids suspended in water. It is assumed that there are no fundamental differences between the principles involved in the treatment of coal slurries and the phenomena encountered in general mineral-dressing practice, and the latter are-presented in some detail. Basic principles are correlated with present practice for both the anthracite and the bituminous coal fields, and methods are discussed which are designed to utilize best all the material subjected to treatment and suggest a sound approach to the solution of problems relating to the return of clear water to process or to streams. In order to approach the practical phases of the subject with sufficient understanding of the basic principles, it is necessary to discuss two types of phenomena that are applicable. The first is flocculation and the second is sedimentation. Descriptive matter is given covering the various methods and types of equipment employed to conduct these operations, with emphasis on their application to water and coal-slurry treatment. DEFINITION OF TERMS Definitions of terms are consistent with modern developments, and conform, in general, with their usage in mineral-dressing practice. Because of variations in coal-preparation methods, it is impossible to define a rigid upper limit of particle size for a coal slurry or sludge which will apply generally. Limits vary somewhat with the particular treatment method employed at a given plant. For purposes of definition, this limit is chosen as the particle size for which preparation methods such as Chance comes and jigs are usually dispensed with, and other methods, such as the Hydro-tator, tables, and froth flotation, are employed. Generally the limit will be 10 to 6 mesh.

Jan 1, 1950

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300' F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300' F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By H. W. Hitzrot

THE new 4800-ton fine-grinding plant at the Hollinger Consolidated Gold Mines property at Timmins, in the Porcupine district of northern Ontario, went into operation in November 1937, and represents the culmination of an experimental program. The plant is a moderately low rectangular building of hollow tile on a steel frame, large enough to house four 7 by 15-ft. ball mills, with a 12 by 27-ft. classifier beside each mill. At one end is space for the addition of another mill and classifier or a mill and classifier for the regrinding of table concentrates. This regrinding is now done in the old grinding department of the main mill building. Also within the new building are the four 6-in. sand pumps for lifting the product from the low-dis- charge ball mills up to the classifier feed launders and the containers and appurtenances of a complete dust-collecting system. Together with the usual overhead crane, these constitute the major units of equipment within the new plant. The location is at the lower end of the main mill buildings toward the old tailings pond, now so neatly leveled off and sodded to rival the very best of our own W.P.A. projects. Close by the new grinding plant is a large dome-shaped structure-the new fine-ore bin designed arid built by Mr. Longmore and the Hollinger milling and engineering staffs. There are four separate automatically loaded inclined belt conveyors leading up from the bottom of the bin to the grinding plant. Each belt is equipped with a recording weightometer, in the grinding building. The whole structure and the machinery within it give the impression of permanence and a carefully integrated design with close attention to details, giving the operators every advantage for consistently good work. The many ideas and opinions behind this achievement were all ably organized and executed by the Hollinger milling and engineering staffs, anti not a few were directly initiated by them. Excellent cooperation was afforded by the Hollinger staff to the engineers helping to conduct the grinding and classification test program. It was the author's privilege

Jan 1, 1939