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DAVID COLE, El Paso, Tex. (communication to the Secretary*).-I notice that Mr. Callow takes issue with me on certain points I have previously contributed to the Transactions, to which I would reply," follows: Regarding Inspiration: Mr. Callow's refinements of gravity methods made just prior to the advent of flotation in the Southwest were intended to be "down to the minute," and the gravity flow sheet for Inspiration ores which he brought out at that time as the result of his long campaign of experimentation was regarded by many as the last word in wet gravity work. It was not my intention to discredit this good work when I said "he threw small light on slime treatment." "Minus 200 mesh," however, is not-synonymous with "real slime." Repeated treatments by gravity methods will result in the separation of a very large percentage of the fine granular sulphides in sizes much finer than 200 mesh, and I gave him due credit for that. The unavoidable losses were made in the "real" slime which I mentioned in the paragraph which Mr. Callow refers to as my "statements on page 1871," and his refinements were not a solution to the real slime problem. Mr. Callow has called attention to the fact that his best results at. Inspiration in gravity work were obtained on Joe Bush ore and he presents a table, viz., Tests Nos. 68, 71, and 72; and asks us to consider and compare these results, made by his refined gravity methods, with the results obtained by 'flotation methods in the new mill "during the second half of 1915 and in August, 1916." Of course, the ore treated in the mill was what happened to come. It was not Joe Bush ore, and therefore comparisons of detail figures are of no value. In regard to Mr. Callow's statements under the caption "Callow Cells Adopted by Arizona Copper Company," I would say that those who have been sufficiently interested to read my paper and' his remarks will notice what amounts to contradictory statements on several points which I cannot allow to go unchallenged. I was on the ground constantly and got my information at first hand. The tables published were the official ones. Mr. Callow was not present during the period and has been laboring under that disadvantage.

Jan 5, 1917

By F. V. Lenel, G. S. Ansell, E. C. Nelson

IRMANN'S' discovery that extrusions of fine alumi-num-flake powders possess remarkable high-temperature strengths has led to the production of a new class of engineering materials whose properties are derived from a fine dispersion of oxide particles in a metallic matrix. A fundamental study of such materials and elucidation of the mechanism whereby the oxide parti-

Jan 1, 1958

By R. W. Nash, A. E. Schwaneke, V. R. Miller

The Bureau of Mines developed a detector for controlling sorting devices to separate the copper-bearing fragments from the barren portion of Michigan native copper and western prophyry copper ores. A successful sorter could be used to preconcentrate the ore and eliminate costly hoisting or hauling of waste. An induction balance unit was developed and used in sorting tests on crushed sized - 2.5 + 1.3-cm (- 1 + 112- in.) ores from the Kingston (0.9% Cu) and Centennial (1.6% Cu) mines of Upper Michigan. Concentrates of 4.0% and 7.5% Cu, respectively, were obtained at recoveries of 85 and 80% of the copper in the feed. Eighty percent of the sorted ore was rejected as tailing. Sorting larger fragments improved the separation efficiency. Tests on - 10+ 5-cm (- 4 +2-in.) sized Kingston (1.6% Cu) and Centennial (2.1 % Cu) ore samples yielded concentrates of 4.3 and 4.5% Cu, respectively, with corresponding recoveries of 92 and 91 %. A bout 60% of the sorted ore was discarded as tailing. Porphyry copper sulfide ores also responded to electronic sorting but at a lower separation efficiency. Sorting response was affected by the degree of segregation.

Jan 1, 1979

By A. Rohkohl, R. Piillenberg

The QSL plant at "BERZELIUS" Stolberg GmbH in Stolberg, Germany was commissioned in 1990, introducing a new era of lead smelting in the more than 150-year-old history of the lead smelter. The lead smelter in Stolberg has changed its appearance several times during its existence. Therefore, at many locations on the site, old and obsolete buildings and equipment parts can still be seen. They are often located side by side with new modern constructions like the QSL-plant and the mercury removal plant, and reflect the specific contrast between tradition and modern technology. After a brief description of the development of the smelter and its decision to implement the current novel plant, the most recent operating results are reported. Following some modifications to the process and equipment, which had become necessary along with the implementation of the new technology, the plant is in operation today at approximately 30% above its design capacity. These achieved production levels are not restricted by the QSL-reactor but by precedent and subsequent available equipment. The QSL-plant entirely complies with the demands of modern, energy-saving, ecologically compatible lead production technology. Stringent environmental standards are maintained on a continuing basis. The demonstrated viability and flexibility of the overall process permits the operation to adapt quickly to the prevailing economical situation worldwide.

Jan 1, 2000

By Visnovsky

The computing needs of a mine planning office vary greatly depending on the staff available and the nature of the work undertaken. Some offices work acceptably with few aids apart from hand held calculators while others use mainframe computers with associated data bases and sophisticated mine planning software. Not every mine planning office requires the big system approach but all can benefit from the intelligent use of small computers. There is virtually no limit to the design problems which can be solved using modern desktop computers. The simple language required encourages engineers to work interactively to facilitate both large and small computations. However a limit does exist in the speed and ease with which solutions can be obtained. The important ingredients for the successful application of such machines are a little computing knowledge, a flexible machine, and an engineer who really understands the mining problems confronting him. If these are available it is possible for the engineer to tailor the program to the task at hand and efficiently achieve the required result.

Jan 1, 1983

By W. M. Dressel

Over 2 million tons of steel furnace dust containing approximately 100,000 tons of zinc and 10,000 tons of lead are produced each year. Recycling of the dusts would conserve mineral resources, lessen land area requirements for disposal, decrease pollution problems, and aid in solving our balance of payment problem. Steelmaking dusts can be pelletized and thermally treated to remove lead and zinc and to reduce the iron oxide to form metallized pellets. In laboratory tests more than 99 pct of the zinc and 85 pct of the lead was volatilized from the dust by roasting at 1,000° C in carbon monoxide. The volatilized material can be collected and treated further for lead and zinc recovery. Reduction and volatilization of over 95 pct of the zinc and most of the lead was achieved by roasting self-reducing mixtures of steel furnace dusts and high carbon steelmaking wastes in an inert atmosphere at 1,100° C. Over 80 pct of the iron in the resulting pellets was metallized.

Jan 1, 1974

By Stille H

The choice between drill and blast and mechanical boring methods is often not obvious. There are several factors which may have a de- cisive influence on this choice, such as the geometry and length of the tunnel, site con- ditions and borability of the rock. Both the un- confined compression strength and the joint dis- tance have a considerable influence on the net penetration rate. Site conditions, rock quality and water conditions influence the utilization of the tunnel boring machine. The drill and blast method is normally less sensitive to devi- ations from assumed ground conditions.

Jan 1, 1987

By V. R. Garfias

. EVEN before the arrival of Cortes in 1519, the history of Mexico was closely linked to that of its mineral production; the mining activities of the Aztecs being thus described by Clavigero, one of the early historians: In the mountains of Anahuac there abound in veins all kinds of metals, and infinite variety of other fossiliferous products. The Aztecs obtained gold from the countries of the Cohuijques, of the Mixtecas, of the Zapotecas and from various other places. They gathered that precious metal usually in grain, from the sands of the rivers, keeping a certain portion for the crown. They obtained silver from the mines of Taxco, of Tzumpango and others; however, this metal was not as esteemed by them as by other neighboring nations. They had two kinds of copper; one hard, which they used instead of iron to make sickles; lances and all kinds of rural and military instru-ments and another, soft, with which they made kettles, goblets and other receptacles. This metal abounds mainly in the province of Zacatula and in that of Cohuijques, which at present is in the kingdom of Michoacan. They obtained tin from the mines of Taxco and lead from those of Izmiquilpan, located in the country of the Otomies. From the tin they made money, and the lead, we know they sold in the market places, but we ignore the uses to which they applied it. They also had mines of iron in Tlaxcala, in Taxco and in other places; but they either had not discovered them, or did not know how to make use of the metal they contained. In Chilapa there were mines of mercury, and in other places there were mines of sulphur, alum, vitriol, cinnabar, cohre and of a white earth which they esteemed highly. As regards mercury and vitriol, we do not know of what use (they were to them; the other metals they used in paintings and dyes. There were then great abundance of amber and asphalt-bitu-men of Judea-on the coast of the seas and many cities of that territory paid tribute of one and the other to the kingdom of Mexico. They mounted the amber in gold, and it was only used as ornament and display by them. With the asphalt they made certain perfumes.

Jan 1, 1937

By L. S. Darken, H. A. Wriedt

In a previous paper,1 an experimental study of nitrogen dissolved in a cold-rolled ad heat-treated, low-carbon steel at 300° to 450°C yielded the equilibrium solubility relations and the concentrations of the two types of deject sites involved. In this paper, the probable identities of the defects that absorb nitrogen are restricted from the available evidence to screw dislocatioms for B sites, and to either micro crack surfaces or edge In a previous paper.' we reported an experimental study of the excess nitrogen solubility in a low-carbon steel produced by a combination of plastic deformation and nonrecrystallizing heat treatment. The concentrations of the defect sites that absorb the excess nitrogen were measured in terms of their capacities for nitrogen. Also measured were the equilibria of nitrogen distribution among the sites. Several problems and questions arise from these measurements. We should like a physical identifi- dislocations for A sites. The concentrations of these defects in experimental steel specimens are estimated from the concentrations of defect sites. finally, the influence of chemical composition upon the nitrogen-trapping defects—the relative numbers generated in deformation, the relative numbers surviving as temperature is raised, the apparent numbers observed, and the distribution equilibria-is discussed. cation of the defects involved and of the two kinds of nitrogen-absorbing sites, called A and B types, upon those defects; the concentrations of the defects themselves could then be computed. We should also like to know how exclusively the findings apply to the "X steel" used in the experiments and to what extent they apply to cold-rolled steel given no subsequent treatment. By considering the data already given, together with results from a few specific tests which will be reported as they are invoked, we are able to eliminate some interpretations altogether and to secure an acceptable though limited interpretation from among the remaining possibilities.

Jan 1, 1965

By George S. Rice

THE A.I.M.E. Ground Movement and Subsidence Committee, pro- posed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bituminous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Pitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O'Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1939

By B. P. Watson

In South Africa, the Bushveld platinum group metal deposits are unique in that they are the largest platinum group metal deposits in the world. These deposits occur as two distinct stratiform tabular orebodies and strike for many hundreds of kilometres. Mining is extensive, with depths ranging from close-to-surface to 2 000 m. The mining method is a variation of planar open stoping. Pillars are widely used to support the open stopes. Crush pillars are commonly used in this role, where the residual pillar strength provides the required support resistance to stabilize the stoping excavations. This paper describes the direct measurement of stress within these crush pillars, and describes the stress?strain behaviour of these pillars. These findings indicate that the measured pillar failure stress is too high. The implications of these findings are that roof damage and pillar bursting could occur, and some examples of this type of damage are shown. The paper concludes that this problem could be mitigated if the three dimensional pillar geometry were modified and the pillars were cut smaller.

Jan 1, 2007

By M. Lahtinen, B. Saxén, M. Latva-Kokko

The conventional jarosite precipitation process with neutralization by zinc calcine is widely utilized in the zinc industry. However, it has a significant drawback of causing losses of valuable metals like zinc, lead and silver to the jarosite. This comes mainly from the content of these metals in the calcine used for neutralization. As the zinc plants today want to maximize metal recoveries and minimize the amount of residues, this matter has become more and more relevant for zinc producers. Outotec developed and patented in early 2000 an improved version of the jarosite precipitation process in which the need for neutralization can be minimized, thus the metal losses to the jarosite also can be minimized. At the same time, the iron content of the jarosite increases and the amount of produced jarosite is smaller. The main feature of this process is to start the precipitation from ferrous iron after having reduced all the iron by zinc concentrate in a separate step before jarosite precipitation. Iron is then oxidised back to the ferric form simultaneously with jarosite precipitation. By this method, the formation of sulphuric acid is only half that compared to the conventional method and the need for calcine for neutralization can be avoided. High mixing and oxygen dispersion capability are needed in this jarosite precipitation process. Such capabilities are found in the Outotec OKTOP® reactor series. High performance reactors are not only needed to generate high quality precipitates, but also to maximize oxygen utilization; i.e., to minimize operation costs. Outotec’s latest development in this area is the OKTOP® plant product in which not only the reactors, but also all the accessories are designed, manufactured and delivered by Outotec. One responsible supplier ensures efficient and fast delivery and can provide mechanical and performance guarantees for the entire plant unit.

Jan 1, 2016

By Benjamin N. Webber

THE peninsula of Nicoya contains virtually all of the known manganese deposits of Costa Rica. These are south and west of the Tempisque River, which flows across the peninsula near its juncture with the mainland. This peninsula forms the westernmost part of the province of Guana¬caste, which is the northern of the two Pacific Coast provinces of Costa Rica. The larger villages are connected by ox-cart roads and trails. The area may be covered on horseback without difficulty except during the rainy season, when rivers are in flood and difficult to ford. This is particularly true of the Rio Seco-Guayjiniquil area, where there are no inhabitants and few open trails through the tropical jungle. Santa Cruz and Sardinal are served by the Taca Airlines, two planes a week from San Jose and Puntarenas. The port of Puntarenas has excellent loading facilities but none are available nearer the manganese districts; there ore must be lightered to vessels standing offshore. PHYSIOGRAPHY In Costa Rica there are three physiographic divisions: (I) the central plateau (Mesa Central), a somewhat dissected lava platform upon which are superimposed volcanic peaks, some of which are still active; (2) the Atlantic lowlands and (3) the Pacific lowlands. The peninsula of Nicoya and the larger part of the province of Guanacaste lie within the Pacific low lands. The term "lowland" indicates relation to the central plateau, as the peninsula of Nicoya is largely mountainous. The mountains of Nicoya form the southern extremity of a range of low hills, which are almost continuous along the Costa Rican coast to the Nicaraguan border. The coast line shows evidence of subsidence. The Gulf of Nicoya obviously is the drowned valley of the Tempisque River. The west coast of Nicoya shows many rocks, shoals, bays, inlets and small islands as evidence of a sea encroachment. The altitude of the manganese deposits ranges from 0 to 800 ft. Topography is of the mature type of wide valleys and low ,hills, often with abrupt slopes. The climate is tropical and humid; rainfall of 125 in. falls largely between Sept. I and Jan. I. Vegetation often is very dense, the Rio Seco-Guayjiniquil area being truly tropical jungle, containing many hardwoods. HISTORY AND PRODUCTION The production of manganese ore in Costa Rica apparently dates from the first World War. It reached a peak of about 10,000 tons in the year 1919, then declined rapidly. Evidence of this period of prospecting has rotted away or been swallowed by the jungle. The war now in progress apparently revived the interest in Costa Rican manganese and has encouraged further prospecting. Deposits long known have been redenounced and trails have been cut through the jungle to them. Many new deposits have been discovered in the last

Jan 1, 1942

By C. E. Capes

INTRODUCTION In the coal industry, ever greater amounts of fine coal must be processed due to the need to crush to finer sizes for impurity liberation in low cost physical cleaning processes. Fine coal processing is a major problem for the industry, both through the cost of disposal of reject fines and through the mining cost of lost coal values. Of the available disposal methods, impoundment has always been the most popular. In many cases, when the coal fines were of sufficient quality, froth flotation and expensive filtration methods were used to recover the coal in these waste streams. The coal fines are not always of a marketable quality, however, and flotation, due to poor yields and difficult dewatering, has always been an expensive and not necessarily profitable process. The oil agglomeration method is increasingly important as a technology for the treatment of fine coal-in-water suspensions. These fines can be preferentially wetted and agglomerated by many different oils mixed with the suspensions. Ash-forming impurities remain in suspension and are rejected when the agglomerates are recovered, for example, by screening. Agglomeration-based methods provide perhaps the only practical means for upgrading extremely fine coal to useful products on a large scale. As with froth flotation, oil agglomeration relies on differences in the surface properties of coal and impurities to effect separation. Whereas flotation be- comes less effective with particles finer than 200 mesh (0.075 mm) or coarser than 50 mesh (0.3 mm), there appears to be no lower limit on particle size for oil agglomeration while particles % in. (3.175 mm) diameter or larger can also be treated, if desirable. In addition to this ability to clean a broader size range of coal particles, oil agglomeration can produce a dense, coarse granular product of acceptable strength and low water content without the need for vacuum or thermal drying. The preferential wetting of solid surfaces by immiscible liquids has been utilized, if not fully understood, for a long time. Indeed Gaudin,229 in tracing the ancestry of froth flotation, refers to a 15th century oil-pulp kneading procedure in which water-wetted particles were freed under water from an oiled mass. An understanding of the physical chemistry involved in such

Jan 1, 1991

By H. J. Stehli

The first sintering of zinc ores was done at the Bartlesville plant of the National Zinc Co., under the auspices of the late Otto Rissman. Mr. Rissman, who had had a long experience in the treatment of zinc ores, well realized all the difficulties of satisfactorily roasting zinc concentrates, which were becoming finer and dustier as each year went by—thus increasing the difficulties of roasting them in the existing forms of roasters, which were furnaces of the Ropp, Hegeler or similar types. Mr. Rissman's idea of the advantages of sinter was expressed in this way: in a zinc ore dead-roasted in a mechanical furnace there are many particles that are just like an egg; that is, they have a hard and impervious film of fused material on the outside, and in the retorting this completely cuts off access of the reducing gases to the center of the egg until a temperature has been reached at which this shell of the egg can be broken down through fusion. A particle of sintered zinc sulphide, on the contrary, is exactly like a piece of sponge. The moment the retort reaches the temperature at which the reactions can take place, they begin, because the reducing gases have ready access to the interior of the particles by means of the pores of the '" sponge." His faith in the advantages of sinter was amply borne out by the results obtained at the pioneer sintering plant for zinc ores installed by the National Zinc Company. Early Method of Sintering The method of procedure at Bartlesville was to roast the zinc concentrates down to somewhere in the neighborhood of 8 per cent residual sulphide sulphur; to cool these, then to wet them down to the proper moisture and mix them in the mixer of the sintering plant; then to sinter them by means of the heat available from the residual sulphides. The sulphur contents of the roast varied considerably, of course, as it was difficult with the furnaces in use to control the roasting exactly at any point short of dead roast. Hence at times the concentrates were roasted a little too low to sinter readily, and at such times the lack of sulphur was made up by the addition of raw concentrates to the sintering-machine charge in suitable amount.

Jan 1, 1937

By James R. Brown

Mount Isa Mine should be regarded as a large mining complex rather than one mine. This arises from the geography, variety and extent of the orebodies, the mining methods, and the use of 14 shafts for ventilation distribution. A brief resume is presented here since comprehensive descriptions of lead-zinc mining methods {11}, copper mining methods {2 }{3}, geology 241, stope fill {5} and ventilation techniques {6}{7}, have all been published in the last five years. Mining extends over a north-south strike length of 2.5 miles (4 Ian) with a maximum width across dip approaching 3 200 ft (1 Ian). Current mining is to a depth of 3 150 ft (960 m) with the ore hoisting shafts extending 660 ft (200 m) deeper. The composite longitudinal section, Figure 1, shows a representational disposition of the copper and lead-zinc-silver orebodies in the central shaft complex where the main service and hoisting shafts are located. To the south the X41 man and materials shaft provides access to the copper area. The mine layout is typical of high tonnage track- less mines, utilising a large diesel fleet, raise- borers and some large diameter hole stope drilling. Mining methods include sublevel open stoping, slot and massfire, and a trial vertical crater retreat in the copper ore, and sublevel caving, sublevel open stopes with and without rib pillar recovery, and mechanised cut and fill in the numerous lead orebodies. All ore reports to one of the three crushing systems located below 19 level in the central shaft complex. Copper ore is hauled exclusively along the 19 level computer assisted Rockflo haulage system, and lead-zinc-silver ores report via central orepasses from Granby haulages on 13 level and 15 level. Tonnages hoisted during the 1977178 financial year were 5.0 million for copper at 3.2%. and 2.3 million for lead-zinc-silver at 7.8%. 6.3% and 7.06 oz/ton (216 g/tonne) respectively.

Jan 1, 1980

By E. A. Anderson

IN another paper1 the writer has shown that the low permissible maximum fiber stress in a loaded zinc sheet demands a much closer support spacing than is used for steel. The limiting fiber stress in 13-gage (0.032 in.) sheets having a 1-in. depth of corrugation was found to be about 5000 lb. per sq. in. on a 37-in. span under a load of 40 lb. per sq. ft. Corrugated steel sheets are commonly laid with spans varying from 48 to 60 in. In order to increase the span in 13-gage zinc to 54 in. (a common spacing) the metal must be capable of withstanding, under long-continued loading, a maximum fiber stress of about 10,000 lb. per sq. in. This is known to be impossible with unalloyed zinc. In this paper the writer describes a series of experiments leading up to the development of an alloy capable of safely sustaining this fiber stress. The effect of alloying on the slow creep of zinc was studied by means of static tensile tests, accurate determinations of apparent elastic limits and finally by means of loading tests on full-size roof sections. Two industrial buildings roofed with alloyed zinc sheets on spans involving fiber stresses in excess of 10,000 lb. per sq. in. have been under observation for over two years without signs of sagging.

Jan 1, 1929

By W. C. Ellis

Growth directions of tin, zinc, and cadmiun whisker crystals grown spontaneously fi-om the solid are summarized. In tin the edge and diagonal directions of the unit cell are jound with the exception oj- the [110] diagonal. In cadmiurn and zinc the usual dip-ections are a or a + c. The c direction has not been ohserwed. The gvouth directions from the solid are for the most part the same as those found for vapor-grozun crystals and are the ohserz:ed glide directions in plastic j-lou. Whisker crystals of these elements grown from the solid are not twinned. Specular lateral reflections are associated with planes of snall crystallographic indices. WHISKER crystals of tin, zinc, or cadmium grow spontaneously at room temperature in periods from months to years from thin electroplates and other prepared films of the respective metals.''2 It has been established that these crystals are single, but the question of twinning, such as occurs in whisker crystals of silicon,3'4 has not been resolved. Studies of morphology have been restricted to determinations of growth directions.5-' Whisker crystals are both straight and kinked. A few studies have been made on kinked whiskers including measurements of kink anglesa and determinations of growth directions,' but unfortunately not both on the same crystals. The kink angles, for the most part, for each species are grouped closely about a few unique values, but these values are not always those between observed growth directions in a single crystal. To supplement existing data, the growth directions, identification of lateral features, and presence or absence of twinning were determined with a Weissenberg goniometer by a method described elsewhereg on eight whisker crystals of tin, five of zinc, and five of cadmium. One kinked cadmium whisker was examined in detail. The specimens were needlelike, a few microns in cross section, and of the order of a millimeter in length. They were removed from electroplated substrates on laboratory specimens and commercial equipment after several years of growth as described in Table I. Since the specimens were less than 10 p in cross section they were ideally strong with no attendant risk of plastic deformation or damage in handling. Only deformation twinning has been reported for these elements. Therefore, twinning probably should not be expected in whisker crystals of these metals, particularly a twin plane containing the growth direction. A twin plane oriented in this direction was the one sought in this study. No examination was made for twinning at the boundary between kink segments except for one cadmium crystal. Well-defined reflections occur as a whisker crystal of tin, zinc, or cadmium is rotated about its growth direction when illuminated by a beam of light Although these reflections had been noted in earlier investigations, no study of them was made, probably because the earlier studies had disclosed also that

Jan 1, 1967

By Fred C. Bond

The process of comminution by grinding is properly classified as an art, rather than as a science. Like most other operations concerned in ore dressing, or in the treatment of nonmetallic minerals, the mathematical relationships of rock breaking are still obscure, and much disputed. If some basic formula exists, comparable for instance to E = RI in electricity, it has not yet been placed in common use. The absence of such a comprehensive mathematical background undoubtedly is the cause of many inefficient designs and installations, as well as of much poorly oriented experimental work. On the other hand, the mathematical immaturity of the process constitutes an opportunity for a real contribution on the part of those who enjoy mental pioneering. Some of the statements made in any such exploratory work may be little more than opinions, and it is to be expected that some will later be proved untrue; but such an evolutionary development is necessary if we are ever to arrive at the fundamental laws, the discovery of which can change the empirical art into a mathematically exact science. The most significant principle of comminution that we possess is probably Rittinger's law, which states that the work done in crushing or grinding is proportional to the area of the new surface produced. There is an alternative theory, known as Kick's law, which formerly was preferred to Rittinger's but which now is rather discredited. According to Kick, the work done in comminution is proportional to the volume, or weight, of the pieces broken, or to the cube of their diameters, while according to Rittinger it is proportional to the square of the diameters. Careful determinations of size distribution below the limit of sieve sizes1 indicate strongly that under ordinary grinding conditions the amount of new surface area produced is directly proportional to the net power input. Changes in the size ratio between the balls and the particles cause important differences in the proportion of the product that passes a given mesh size, but do not affect the total amount of new surface area; they merely change the distribution of the surface among mesh sizes. The principle of constant surface area is the logical result of Rittinger's law, and it may well have an important bearing on the comminution theories of the future. Formula for Ball Wear Grinding tests have been made in the Allis-Chalmers laboratory for more than 9 years, using a standard charge of mixed balls of different sizes; and a continuous record has been kept of the rates of wear of each size. These results have been combined and plotted in various ways in order to discover the relationship that exists

Jan 1, 1943

By J. R. Lewis, J. H. Hamilton, V. D. Hanson, K. W. Heinlein, J. M.194-000-000-006 Vonfeld, J. P. Denny

TWO overfeed drum-type separators using a suspension of magnetite in water as the separating medium have been installed in the Champion NO. 1 preparation plant of the Pittsburgh Coal Co., Division of the Pittsburgh Consolidation Coal Company. A simplified flow diagram of the Champion No. 1 coarse coal circuit is shown in Fig. 1. A railroad car rotary dump, capable of handling 20 cars per hr, discharges the coal into a hopper. Coal is fed from the hopper of the rotary dump to a 60-in. belt conveyor by two reciprocating pan-type feeders. The belt conveyor feeds the run of mine coal onto the main raw coal screen. The first deck is fitted with 3 in. round hole perforated plate which removes the 3x0-in. coal. The second deck of the screen is equipped with 6-in. lip screens which separate the remaining coal into +6 in. and 6x3 in. The +6 in. coal is transferred to a shaking picking table and then discharged onto the loading boom. The 6x3-in. coal is fed to the No. 2 dense medium vessel for cleaning. The clean coal from this unit flows to a vibrating screen where the medium is removed and the coal sized into 6x4 in. and 4x3 in. The 6x4-in. coal is transferred to the loading booms and the 4x3-in. is conveyed to the clean coal classifying screens to remove degradation. The 3x0-in. raw coal is blended in a system composed of three storage bins having a total capacity of 400 tons. From these bins the coal is fed onto a 48-in. flight conveyor by three reciprocating feeders. This flight conveyor delivers the coal to the Rheo plant. The 3x0-in. coal is then cleaned in two 48-in. Rheolaveur launders each equipped with two Rheo boxes. The first set of boxes removes a primary refuse and the second set produces a middling product which is recirculated in the launders. The cleaned coal and water overflows from the primary launders to two sizing and dewatering shaker screens where the water and — -in. coal are removed for further cleaning in the Rheo fine coal plant. The 3x3h-in. clean coal is combined with the 4x3-in. coal from the No. 2 vessel, sized into 4x2 in., 2x1 34 in., and 1'/4x in., then sent to the loading booms. The 3x0-in. refuse from the primary launders is normally fed to the No. 1 dense medium circuit, but any portion or all of this material can, as an alternate, be processed in the Rheo rewash launder. The primary refuse when fed to the dense medium circuit is screened into YSXO which is cleaned in the fine coal plant and 3x for cleaning in the No. 1 vessel. The clean coal from this vessel is passed over a vibrating drainage and rinse screen, then sluiced to the rewash sizing shaker where it is dewatered and sized for loading. The installation of a pilot plant separating vessel using a suspension of magnetite in water was pro-

Jan 1, 1953