Search Documents

By J. C. McCarthy, J. B. Stone

AT Fresnillo a series of veins that has yielded very large quantities of silver and other metals has been developed over a length of 6500 ft. and to a depth of over 3000 ft. In the course of this work striking changes in the mineralogy and metal content of the principal veins have been disclosed and are well illustrated by assay data and ore reserve calculations. This paper describes those changes and suggests their bearing on theories of ore deposition. LOCATION The Fresnillo mines and the adjacent city of Fresnillo are in the central part of the state of Zacatecas, 35 miles northwest of the capitol city of Zacatecas, once a great silver-mining center whose total production is said to place it fourth in the world. Another famous old silver district, Sombrerete, lies northwest of Fresnillo. Fresnillo is connected by a 5-mile branch with Fresnillo station on the Mexican National Railways line from Juarez to Mexico City. Lying near the southwestern edge of the central plateau of Mexico, Fresnillo ís surrounded by a semi-arid rolling plain from which rise widely scattered low hills. One of these is Proaño hill, an inconspicuous elevation rising barely 300 ft. above the plain, but in and around Proano hill are the Fresnillo mines, and it is marked now by a great gash cut by glory-hole mining. HISTORY According to an account written in 1834 by Francisco García, then Governor of the state of Zacatecas, Fresnillo was discovered in 1554, and a presidio was established there in 1568 to protect travelers on the road between Zacatecas and Sombrerete. Little is known of the earliest mining at Fresnillo, which seemingly began soon after its discovery, but by piecing together various accounts, it appears that mines there were working in 1717 and were notably active ín 1751 under an administrator named Murguía, who left detailed notes describing them. In 1757, however, the mines became inactive because of difficulties with creditors and no doubt also because of increasing water troubles, and they remained abandoned until 1830. In 1830 the mines at Fresníllo became the property of the State of Zacatecas because the interest of Governor García had been aroused by the long disregarded notes of Murguía. Under García's energetic leadership the mines were reopened, and silver production began in 1832 ín spite of serious difficulties, including abnormal rains and an epidemic of cholera. In 1835 English capital was sought to allow installation of steam-driven Cornish pumps at the two principal shafts and to build the great hacienda de beneficio (the present Proaño hacienda). The company formed in 11835 (Cia. Zacatecano Mexicana) operated successfully until 1872, when its career was brought to an end by the political and economic disturbances

Jan 1, 1942

By J. C. McCarthy, J. B. Stone

AT Fresnillo a series of veins that has yielded very large quantities of silver and other metals has been developed over a length of 6500 ft. and to a depth of over 3000 ft. In the course of this work striking changes in the mineralogy and metal content of the principal veins have been disclosed and are well illustrated by assay data and ore reserve calculations. This paper describes those changes and suggests their bearing on theories of ore deposition. LOCATION The Fresnillo mines and the adjacent city of Fresnillo are in the central part of the state of Zacatecas, 35 miles northwest of the capitol city of Zacatecas, once a great silver-mining center whose total production is said to place it fourth in the world. Another famous old silver district, Sombrerete, lies northwest of Fresnillo. Fresnillo is connected by a 5-mile branch with Fresnillo station on the Mexican National Railways line from Juarez to Mexico City. Lying near the southwestern edge of the central plateau of Mexico, Fresnillo is surrounded by a semi-arid rolling plain from which rise widely scattered low hills. One of these is Proano hill, an inconspicuous elevation rising barely 300 ft. above the plain, but in and around Proa[n]o hill are the Fresnillo mines, and it is marked now by a great gash cut by gloryhole mining. HISTORY According to an account written in 1834 by Francisco García, then Governor of the state of Zacatecas, Fresnillo was discovered in 1554, and a presidio was established there in 1568 to protect travelers on the road between Zacatecas and Sombrerete. Little is known of the earliest mining at Fresnillo. which seemingly began soon after its discovery, but by piecing together various accounts, it appears that mines there were working in 1717 and were notable active in 1751 under an administrator named Murguía, who left detailed notes describing them. In 1757, however, the mines became inactive because of difficulties with creditors and no doubt also because of increasing water troubles, and they remained abandoned until 1830. In 1830 the mines at Fresnillo became the property of the State of Zacatecas because the interest of Governor García had been aroused by the long disregarded notes of Murguía. Under Garcia's energetic leadership the mines were reopened, and silver production began in 1832 in spite of serious difficulties, including abnormal rains and an epidemic of cholera. In 1833 English capital was sought to allow installation of steam-driven Cornish pumps at the two principal shafts and to build the great hacienda de benefacio (the present Proano hacienda). The company formed in 1835 (Cia. Zacatecano-Mexicana) operated successfully until 1872, when its career was brought to an end by the political and economic disturbances

Jan 1, 1942

By Jury P. Nazarov

"At present the traditional study of crystal lattices of nmerials in a topic “flota tion mineralogy ” is used up. The present investigation is development of the a sclsical approaches to t he analy sis of ideal crystal lattices and f ormation on the basic obtained results of consultation, having practical chara cter.The strategy of develop ment floatation reagents i s offered, which basic pr inciple – complementary assembling m olecules forming surface co mpounds so thatth e steric charact eristic of the “visitor "" (of a polar gro up of molecule SAS) corresponded to the st er ic requirements of the ""o wner"" (elementary cell of a mineral), or coincided inter molecular cavities). In itsh case process of for mation of surface com pounds ca n be considered(a first approximation) as a realization of a principle of the design er "" LEGO "".The approach, which basic idea consi deration actioonf reactant ions and ions of ele mentary cell as three-di mensional objects is developed. And at suchp paroach the laws of a ster eometry - laws describin g spatial effect s are applicable. Thus it is possible toconsider a miner al surface as a mat r ix - basis for formation of the requirem ents to polar group of a molleec cuollector, and or (it is possible) depressant. As a whole approach represents a generalized principle double com plementation, enveloping both power, and geometric asp ects (concept ""lock "" and “key "" - spatial conform ity. Figure 7 (left). Fragment of a crystal lat tice of ap atite. Oleic acid ions interact on ly with th e calcium cations locate d on a long axi s of an elementary c ell. Magenta colour represe nts the calciu m cations, which are n ot particip atin g in interactio n, as th e degree of screening by ions neighbours exceeds 1.Figure 2. (right)The illustration of space complementary. “Afterbuilding” an elementary cell of apatite (radical PO4) related molecules of the visitor – alkyl phosphate (polar group PO4). Cations of calcium are not shown."

Jan 1, 2003

By A. Prokopiev

The biological stage of reclamation is the third final stage in remediation the damaged mining terrains, which is carried out following the completion of the mining technical stage. The biological stage of reclamation is carried out through restoration of the vegetation layer. The biological reclamation includes all activities related to the introduction into the terrain of vegetation, microorganisms, insects, and animals, with the acceleration of the succession and creating conditions for a sustainable ecosystem close to the natural one. It comprises of all anti erosion and other activities such as the preparation of the land for growing agricultural crops, sowing grass seeds, planting seedlings, amelioration of the soil and further care for the vegetation until its final place on the land. Mineral and organic-mineral ameliorants for biological reclamation were studied. The basic mineral ameliorant is Bulgarian natural zeolites. The zeolitic-rich tuff samples provided from Golobradovo deposit was characterized by their chemical and mineralogical composition and ion exchange properties and X ray analysis. The average content of clinoptilolite was found to be more than 80%. The materials had ion exchange capacity (NH4+) varying from 100 to 110 meq/100g. Chemical modifications of natural clinoptilolites were made in order to form active phases of nitrogen, phosphorus, potassium, and microelements. Slow-acting mineral ameliorant was developed. Organic-zeolite substrate for biological reclamation was developed. This new product is a mechanical mixture of organo-matter and modified zeolites. Agro-chemical parameters of new ameliorants were determined. The substrates including studied overburden dump soil and mineral or organo-mineral ameliorants are produced. The grass growth test was performed. Keywords: biological reclamation, ameliorants, damaged mining terrains

Sep 1, 2012

By J. S. Limb

Actual issues which arose during the discovery process or subsequent court testimony are presented. While most of these issues relate to the author’s specialty of mineral appraisal some are more general in nature and could be of interest to those who may be involved in current or future court testimony. Several important areas to be reviewed and agreed upon with the client and their legal counsel are discussed. General observations on litigation, the discovery process, testimony and some repercussion’s are also given.

Jan 1, 2001

By John T. Neubert

Between 1988 and 1990, the U.S. Bureau of Mines studied the mineral resources of San Juan National Forest to appraise and determine types and locations of deposits that could be mined economically now and in the foreseeable future. The study included a comprehensive field examination and literature search. Substantial amounts of lead, zinc, copper, gold, and silver have been mined from within the Forest in the La Plata, Rico, Silverton, and Dunton areas. At least one buried 200 million ton stockwork molybdenum deposit occurs in the Rico area. Similar size porphyry copper deposits occur at the surface and about 1,700 feet below the surface in the Bedrock Creek area of the La Plata district. Numerous smaller base-and precious-metal and uranium deposits have been prospected and mined throughout the Forest. Coal, oil, and gas have been or are currently being produced in the western and southern parts of the Forest. Carbon dioxide occurs in the western part of the Forest. Strategic elements occurring within the Forest include platinum, palladium, cadmium, gallium, indium, bismuth, and tellurium; however none of these elements is economically recoverable at present. Several geothermal areas are scattered across the eastern two-thirds of the Forest. Industrial minerals within the Forest include sand and gravel, limestone, and decorative stone.

Jan 1, 1992

By D. C. Bard

THESE notes are based on the megascopic study of a suite of 2,400 specimens of minerals and ores from the Butte mines, combined with field observations at intervals over a period of several years. The estimate of average vein composition is based on field acquaintance with the veins. It is probably correct in its broader proportions, and should be of use for comparison. Chemical and Mineral Composition of Butte Sulphide Vein. Matter. Per Cent. Oxygen 43.00-Quartz, rhodonite, rhodochrosite, barite, calcite, and hübnerite. Silicon 36.00-Quartz, rhodonite, and willemite (very rare). Sulphur 5.00-Sulphides of. iron, zinc, copper, lead,. and sulphates of barium and rarely calcium. Iron 5.00-Ferric sulphides (pyrite, bornite, chalcopyrite), also in sphalerite, tetrahedrite, and rarely hematite and siderite. Manganese, . . . . 2.00-Rhodonite, rhodochrosite, and hübnerite. Zinc, . . . . 1.00-Spha erite, very rare wurtzite and willemite, and tetrahedrite and tennantite.

Jan 8, 1913

By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By G. G. Gunther, C. L. Boeke

DISCUSSION I. M. Symonds, (Cia. Minera de Penoles, Monterrey, N. L., Mexico)—Messrs. Boeke and Gunther have written most interesting papers describing their zinc de-leading operations. Mr. Gunther was kind enough to show me his operation last fall. Aside from the excellent results, I was most impressed by the ease of control. I had expected a difficult operation requiring careful control and supervision, but found one requiring little attention. The only disappointing fact of the separation is the poor copper recovery in the concentrate of the delead-ing float. The final zinc concentrate still has about 3 pct chalcopyrite, assuming the copper is chalcopyrite, and adversely affects the zinc grade. If the copper mineral is free, the poor recovery is probably due to the high cyanide concentration. Fortunately, about 75 pct of the copper is recovered before the pulp goes to the zinc circuit. Mr. Gunther's remarks about proper cyanide to zinc sulphate ratios cannot be overemphasized. Although our experience has been confined mostly to the conventional lead-zinc separation, we have found these ratios most important, and more important in the mill than in the laboratory. The laboratory appears to give optimum results over a wider range of reagent control than the mill, especially in regard to the effect of cyanide. In handling relatively high grade ores in the mill we usually find that ratios of cyanide to zinc sulphate of 1 to 10 or 20 or higher work best. Where there is a substantial amount of chalcopyrite present it has been noted that cyanide in the range of 0.02 to 0.10 lb per ton of ore is often beneficial in zinc rejection and possibly in copper recovery in the lead concentrate, while larger amounts, say 0.20 to 0.25 lb per ton, will float more zinc and cause a serious rejection of chalcopyrite. On an ore with 15 to 20 pct zinc content we have had to use 0.8 to 1.2 pounds of zinc sulphate per ton of ore with the above cyanide to obtain optimum results. Some time ago we made a laboratory study following Parral and Santa Barbara practice on an ore which had never a very satisfactory rejection of zinc in the lead concentrate. The results were not encouraging and a mill test was not run. Since then the study has been reopened using in the laboratory fresh zinc concentrate from the mill instead of one made in the laboratory. A reasonably clean lead float from the zinc concentrate was obtained. It has been checked several times and each time the zinc concentrate made in the mill responds and the one made in the labora-

Jan 1, 1953

By R. C. Edwards, T. G. Kirkland, Will Mitchell

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By R. C. Edwards, Will Mitchell, T. G. Kirkland

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By Strathmore R. B. Cooke, Thomas E. Bam

The paper discusses the mineralogy, structure, and strength of magnetite pellets fired in air and in a neutral atmosphere at various temperatures. The information obtained from this investigation is used to explain the origin of the structures occurring in pellets produced by pilot-plant operation. MUCH experimental work has been performed on i-VL the mechanism of agglomeration and pelletiz-ing of moist ore and concentrates and on the process of firing agglomerates, but except for the work of Tigerschiold and Ilmonil much of the published information is empirical, and the literature contains no information concerning the structure of fired pellets. Results of a preliminary study of the micro-structures occurring in pellets produced both in the laboratory and on a pilot-plant scale are presented in this paper. Most of the results were obtained from pellets prepared from Mesabi magnetic taconite, but through the courtesy of the staffs of the Minnesota Mines Experiment Station and the Bethlehem Steel Co., fired pellets produced from limonite and hematite ores and from Lebanon magnetite have also been examined. The work has been restricted almost entirely to pellets produced without additives, the exceptions being those in which coal was added to the filter cake before rolling. Schwartz has described the microstructures occurring in ore sinters produced under reducing conditions.' Many of the structures and constituents he describes are common to fired pellets, but the constituents normally occur in markedly different proportions. His statement that "in all satisfactory sinter the amount of hematite is small, perhaps not over 2 or 3 per cent" is inapplicable to pellets produced by pilot plant operation or by any process in which air is not deliberately excluded. Pilot plant pellets contain over 90 pct by weight of hematite, and those produced in the smaller furnaces of the Minnesota Mines Experiment Station contain more than 80 pct of that constituent. Preparation of Specimens It is difficult to examine unfired pellets microscopically because of their insufficient coherency to withstand sectioning and polishing. Fired pellets sometimes lack coherency and frequently contain holes which are troublesome during polishing. Impregnation of both air-dried and fired pellets is readily accomplished by the use of methyl metha-crylate monomer." The dry pellets are covered in a test tube with excess monomer, to which a small amount of accelerator has been added. When the test tube is placed in a cooling bath and evacuated with an aspirator, most of the entrapped air is removed, and disruption of weak pellets by boiling of the monomer is reduced to a minimum. Polymerization is then effected by immersion of the test tube in a water-bath at about 45"C, this temperature being held for 4 or 5 days. The test tube may then be broken away from the solidified mass and the pellets sectioned by a carborundum wheel or a diamond saw. Where time is important, impregnation is carried out at a higher temperature by Meyer's method.' Air-dried pellets must be impregnated with especial care; otherwise swelling occurs and the solids are displaced. It is more important to be certain that the monomer forms a reasonably complete network throughout the specimen than that all entrained air is removed, and for this reason the vacuum technique may be omitted. Impregnated agglomerated products and pellets are extremely coherent, section well, and may be polished either by hand or by machine. Hand polishing is not satisfactory for the examination of conditions existing at the boundaries of grains, and for this machine polishing must be used. The Graton-Vanderwilt machine gives excellent results, provided that pressure on the specimen is considerably in excess of that recommended by the manufacturer. Mineralogy of Fired Pellets Only two opaque iron minerals, magnetite and hematite, have been recognized in pellets. The complete identity of, these artificial minerals with their corresponding natural counterparts is shown by their mineralographic characteristics and their X-ray structures. Wiistite has not been found. Slag is the most common transparent constituent of fired pellets. Occasionally conditions have been appropriate for the crystallization of fayalite from the slag, and at other times another crystalline sil-

Jan 1, 1953

By Strathmore R. B. Cooke, Thomas E. Bam

The paper discusses the mineralogy, structure, and strength of magnetite pellets fired in air and in a neutral atmosphere at various temperatures. The information obtained from this investigation is used to explain the origin of the structures occurring in pellets produced by pilot-plant operation. MUCH experimental work has been performed on i-VL the mechanism of agglomeration and pelletiz-ing of moist ore and concentrates and on the process of firing agglomerates, but except for the work of Tigerschiold and Ilmonil much of the published information is empirical, and the literature contains no information concerning the structure of fired pellets. Results of a preliminary study of the micro-structures occurring in pellets produced both in the laboratory and on a pilot-plant scale are presented in this paper. Most of the results were obtained from pellets prepared from Mesabi magnetic taconite, but through the courtesy of the staffs of the Minnesota Mines Experiment Station and the Bethlehem Steel Co., fired pellets produced from limonite and hematite ores and from Lebanon magnetite have also been examined. The work has been restricted almost entirely to pellets produced without additives, the exceptions being those in which coal was added to the filter cake before rolling. Schwartz has described the microstructures occurring in ore sinters produced under reducing conditions.' Many of the structures and constituents he describes are common to fired pellets, but the constituents normally occur in markedly different proportions. His statement that "in all satisfactory sinter the amount of hematite is small, perhaps not over 2 or 3 per cent" is inapplicable to pellets produced by pilot plant operation or by any process in which air is not deliberately excluded. Pilot plant pellets contain over 90 pct by weight of hematite, and those produced in the smaller furnaces of the Minnesota Mines Experiment Station contain more than 80 pct of that constituent. Preparation of Specimens It is difficult to examine unfired pellets microscopically because of their insufficient coherency to withstand sectioning and polishing. Fired pellets sometimes lack coherency and frequently contain holes which are troublesome during polishing. Impregnation of both air-dried and fired pellets is readily accomplished by the use of methyl metha-crylate monomer." The dry pellets are covered in a test tube with excess monomer, to which a small amount of accelerator has been added. When the test tube is placed in a cooling bath and evacuated with an aspirator, most of the entrapped air is removed, and disruption of weak pellets by boiling of the monomer is reduced to a minimum. Polymerization is then effected by immersion of the test tube in a water-bath at about 45"C, this temperature being held for 4 or 5 days. The test tube may then be broken away from the solidified mass and the pellets sectioned by a carborundum wheel or a diamond saw. Where time is important, impregnation is carried out at a higher temperature by Meyer's method.' Air-dried pellets must be impregnated with especial care; otherwise swelling occurs and the solids are displaced. It is more important to be certain that the monomer forms a reasonably complete network throughout the specimen than that all entrained air is removed, and for this reason the vacuum technique may be omitted. Impregnated agglomerated products and pellets are extremely coherent, section well, and may be polished either by hand or by machine. Hand polishing is not satisfactory for the examination of conditions existing at the boundaries of grains, and for this machine polishing must be used. The Graton-Vanderwilt machine gives excellent results, provided that pressure on the specimen is considerably in excess of that recommended by the manufacturer. Mineralogy of Fired Pellets Only two opaque iron minerals, magnetite and hematite, have been recognized in pellets. The complete identity of, these artificial minerals with their corresponding natural counterparts is shown by their mineralographic characteristics and their X-ray structures. Wiistite has not been found. Slag is the most common transparent constituent of fired pellets. Occasionally conditions have been appropriate for the crystallization of fayalite from the slag, and at other times another crystalline sil-

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 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 W. Philippoff

ALTHOUGH Gaudin1 and more recently Sutherland2 have calculated the probability of collision of a falling mineral particle with a rising bubble, there is no published information concerning the details of the mechanism of attachment of a collector-coated particle to a bubble. During the past year the writer has developed a theory for the mechanism of attachment, which has been substantiated experimentally. Funds for the investigation and for some of the equipment used have been supplied by the Mines Experiment Station of the University of Minnesota. Motion picture studies of the phenomena involved in the collision between mineral particles and bubbles, such as those of Spedden and Hannan," show that the contact can be completed within 0.3 millisec. Formulas developed for rigid bodies have hitherto been used' for the calculation of the motion of bubble and particle, but it is obvious that a bubble cannot be regarded as a rigid body. On the contrary, Spedden and Hannan's pictures show a great degree of deformation during the collision. The time of attachment was calculated as the time the particle drifted past the bubble. Time of Collision The theory presented in this paper enables calculation of the time of collision, using the concept that the bubble, or more generally, a liquid-air interface, acts as an elastic body. The elasticity, defined as the restoring force on a mechanical deformation, is caused by the surface tension and is the result of the principle of the minimum of free surface energy. It is well known that an elasticity together with a mass determines a frequency of vibration. The vibrations of jets and drops caused by the elasticity of the interface are known to comply exactly with the classical theory of capillarity.' However, the vibrations of isolated bubbles, as distinct from foams, have not been investigated previously. The following equation, presented elsewhere,' has been deduced for these frequencies: 3_____________________ fB = 9.20.vV.vn. (n-1).(n+2)/8 [I] in which fB is the frequency of a harmonic of the bubble in cycles per second, V the volume of the bubble in cc, n a number determining the order of the harmonic, and n = 2 the basic vibration. The first (basic) harmonic describes a change of the spherical bubble to an ellipsoidal bubble. The higher harmonics are more complicated, for the circumference of the bubble is divided approximately into as many parts as the order of the harmonic. As an example, Spedden and Hannan's published motion picture of a vibrating bubble corresponds to the sixth harmonic. Eq 1 shows that only the first and third harmonics are simple multiples (1 and 3), all the others being irrational fractions of the basic frequency. This means that the shape of the vibration can change with time and is in general unsym-metric in respect to the time axis. Such conditions prevail when there is a distributed elasticity or mass, as in the case of vibrating membranes or rods. The constant 9.20 is valid for water at room temperature, but a general solution involving the physical constants of the liquid has not been found. The case of the floating particle is much easier to treat than that of the bubble. It can be assumed that the elasticity is caused exclusively by the interface and that the mass is concentrated in the particle together with some adhering water. The following expression for the frequency of a system of one degree of freedom can be applied: fP = -1/2-vE/m [2] Here fp is the frequency of the particle vibration in cycles per second, E the elasticity in dynes per cm, and m the mass in grams. The classical theory of impact phenomena gives the time of collision during the striking of a spring (in this case the surface of the bubble) by a mass, as: t, = 2/f = nvm/E [3] It is now possible to develop an expression for the elasticity of a floating cylindrical particle. The force equilibrium of a cylinder floating end on at the air-liquid interface is given by the well-known equation (Poisson7 1831) P = /4 D2-pL-g-h + D.y sin a [4] which accounts for the buoyancy and the action of the surface tension where P is the force acting on the particle in dynes (weight-buoyancy), D the diameter of the cylinder in cm, PL the density of the liquid in grams per cc, g the acceleration of gravity = 981 cm per sec2, h the depression of the cylinder below the surface of the liquid in cm, y the surface tension in dynes per cm and a the supporting angles or the one required to insure equilibrium, a being smaller than the contact angle 0. Although demonstrated by Poisson, it has not

Jan 1, 1953

By W. Philippoff

ALTHOUGH Gaudin1 and more recently Sutherland2 have calculated the probability of collision of a falling mineral particle with a rising bubble, there is no published information concerning the details of the mechanism of attachment of a collector-coated particle to a bubble. During the past year the writer has developed a theory for the mechanism of attachment, which has been substantiated experimentally. Funds for the investigation and for some of the equipment used have been supplied by the Mines Experiment Station of the University of Minnesota. Motion picture studies of the phenomena involved in the collision between mineral particles and bubbles, such as those of Spedden and Hannan," show that the contact can be completed within 0.3 millisec. Formulas developed for rigid bodies have hitherto been used' for the calculation of the motion of bubble and particle, but it is obvious that a bubble cannot be regarded as a rigid body. On the contrary, Spedden and Hannan's pictures show a great degree of deformation during the collision. The time of attachment was calculated as the time the particle drifted past the bubble. Time of Collision The theory presented in this paper enables calculation of the time of collision, using the concept that the bubble, or more generally, a liquid-air interface, acts as an elastic body. The elasticity, defined as the restoring force on a mechanical deformation, is caused by the surface tension and is the result of the principle of the minimum of free surface energy. It is well known that an elasticity together with a mass determines a frequency of vibration. The vibrations of jets and drops caused by the elasticity of the interface are known to comply exactly with the classical theory of capillarity.' However, the vibrations of isolated bubbles, as distinct from foams, have not been investigated previously. The following equation, presented elsewhere,' has been deduced for these frequencies: 3_____________________ fB = 9.20.vV.vn. (n-1).(n+2)/8 [I] in which fB is the frequency of a harmonic of the bubble in cycles per second, V the volume of the bubble in cc, n a number determining the order of the harmonic, and n = 2 the basic vibration. The first (basic) harmonic describes a change of the spherical bubble to an ellipsoidal bubble. The higher harmonics are more complicated, for the circumference of the bubble is divided approximately into as many parts as the order of the harmonic. As an example, Spedden and Hannan's published motion picture of a vibrating bubble corresponds to the sixth harmonic. Eq 1 shows that only the first and third harmonics are simple multiples (1 and 3), all the others being irrational fractions of the basic frequency. This means that the shape of the vibration can change with time and is in general unsym-metric in respect to the time axis. Such conditions prevail when there is a distributed elasticity or mass, as in the case of vibrating membranes or rods. The constant 9.20 is valid for water at room temperature, but a general solution involving the physical constants of the liquid has not been found. The case of the floating particle is much easier to treat than that of the bubble. It can be assumed that the elasticity is caused exclusively by the interface and that the mass is concentrated in the particle together with some adhering water. The following expression for the frequency of a system of one degree of freedom can be applied: fP = -1/2-vE/m [2] Here fp is the frequency of the particle vibration in cycles per second, E the elasticity in dynes per cm, and m the mass in grams. The classical theory of impact phenomena gives the time of collision during the striking of a spring (in this case the surface of the bubble) by a mass, as: t, = 2/f = nvm/E [3] It is now possible to develop an expression for the elasticity of a floating cylindrical particle. The force equilibrium of a cylinder floating end on at the air-liquid interface is given by the well-known equation (Poisson7 1831) P = /4 D2-pL-g-h + D.y sin a [4] which accounts for the buoyancy and the action of the surface tension where P is the force acting on the particle in dynes (weight-buoyancy), D the diameter of the cylinder in cm, PL the density of the liquid in grams per cc, g the acceleration of gravity = 981 cm per sec2, h the depression of the cylinder below the surface of the liquid in cm, y the surface tension in dynes per cm and a the supporting angles or the one required to insure equilibrium, a being smaller than the contact angle 0. Although demonstrated by Poisson, it has not

Jan 1, 1953

By Norman F. Schulz, Emert W. Lindroos, Strathmore R. B. Cooke

DURING the course of an investigation of the effects of various starch products on hematite and quartz in regard to their separation by flotation, it was found that whereas most starches flocculated suspensions of hematite in water, they did not flocculate similar suspensions of quartz. However, a derivative of whole corn starch containing approximately one tenth aminoethyl group (-—CH,CH2NH2) per glucose unit proved to be an exception. This aminoethyl starch was a very good flocculant for both hematite and quartz, and also possessed the property of depressing both minerals in cationic flotation tests. In view of the widespread use of starch as a flocculating agent of solids in tailing water, pulps, and for coal slimes, ore and various other purposes in sulphide and nonsulphide flotation, attention is called to the effective surface activity of aminoethyl starch in flocculating suspensions of the two minerals investigated. Test Materials The starch products employed included Globe Pearl starch No. 144 and fresh aminoethyl starch, both supplied by the Corn Products Refining Co. Dr. A. L. Elder" has kindly furnished a description The minerals used in the various tests are listed in Table I. Their specific surfaces were determined by the air permeability method, which gave results reproducible to within plus or minus 3 pct. Surface areas obtained by krypton adsorption were greater than the air permeability values by a factor of 2.0 for quartz and 62.4 for hematite. Starch Adsorption on Mineral Surfaces Starch adsorption data were obtained by agitating for 1 hr measured volumes of neutral starch solutions of known concentrations with weighed quantities of finely divided mineral of measured specific surface. The mineral then was separated from the pulp liquor by centrifugation and the residual starch content determined by chemical analysis. The analytical procedure consisted of oxidizing the starch with an excess of dichromate in acid medium at elevated temperature and titrating the excess dichromate by the ferrous ammonium sulphate—potassium permanganate method. Analytical results were reproducible to within less than 1 pct. Adsorption data obtained in unbuffered solutions near the neutral point are plotted in Fig. 1. The adsorption on either mineral was far greater for the aminoethyl starch than for the Pearl starch. Flocculation The degree of flocculation was determined for a given pulp by measuring the time required for the supernatant liquid to reach a certain degree of clarity. This was accomplished by passing a standard light beam through the pulp, contained in a glass settling tube, and measuring the time required after agitation ceased for the intensity of the transmitted beam to reach a predetermined value. Starch reagent was added to each pulp in small uniform increments, and the settling time was determined after each such addition. Increase in pulp liquor volume due to the added starch reagent was ignored in these tests. Pulps were maintained at or near neutrality. of the preparation of the aminoethyl starch: "Three hundred and sixty grams of powdered unmodified corn starch was slurried in 3 liters of water and stirred during gelatinization. Eighty-six grams of ethylenimine was added slowly during reflux which was continued for 4 hr. The yellow solution set to a gel upon cooling. This was broken up and neutralized with HC1. The thick paste was dried on hot rolls. Adsorbed polyethylenimine and salts were removed from the yellowish products by washing first with acidified aqueous '60 per cent' alcohol. The white product contained nitrogen corresponding to about one aminoethoxy group per ten glucose residues." Only laboratory samples of this starch derivative have been prepared. The starch reagents for use in flocculation, flotation, and adsorption tests were made up by dispersing the appropriate quantities of air-dry starch in cold water, adding boiling water, and heating for 30 min at 120°C under pressure. When used ip adsorption tests, the starch solutions first were cleared of insoluble matter by centrifugation.

Jan 1, 1953

By Strathmore R. B. Cooke, Emert W. Lindroos, Norman F. Schulz

DURING the course of an investigation of the effects of various starch products on hematite and quartz in regard to their separation by flotation, it was found that whereas most starches flocculated suspensions of hematite in water, they did not flocculate similar suspensions of quartz. However, a derivative of whole corn starch containing approximately one tenth aminoethyl group (-—CH,CH2NH2) per glucose unit proved to be an exception. This aminoethyl starch was a very good flocculant for both hematite and quartz, and also possessed the property of depressing both minerals in cationic flotation tests. In view of the widespread use of starch as a flocculating agent of solids in tailing water, pulps, and for coal slimes, ore and various other purposes in sulphide and nonsulphide flotation, attention is called to the effective surface activity of aminoethyl starch in flocculating suspensions of the two minerals investigated. Test Materials The starch products employed included Globe Pearl starch No. 144 and fresh aminoethyl starch, both supplied by the Corn Products Refining Co. Dr. A. L. Elder" has kindly furnished a description The minerals used in the various tests are listed in Table I. Their specific surfaces were determined by the air permeability method, which gave results reproducible to within plus or minus 3 pct. Surface areas obtained by krypton adsorption were greater than the air permeability values by a factor of 2.0 for quartz and 62.4 for hematite. Starch Adsorption on Mineral Surfaces Starch adsorption data were obtained by agitating for 1 hr measured volumes of neutral starch solutions of known concentrations with weighed quantities of finely divided mineral of measured specific surface. The mineral then was separated from the pulp liquor by centrifugation and the residual starch content determined by chemical analysis. The analytical procedure consisted of oxidizing the starch with an excess of dichromate in acid medium at elevated temperature and titrating the excess dichromate by the ferrous ammonium sulphate—potassium permanganate method. Analytical results were reproducible to within less than 1 pct. Adsorption data obtained in unbuffered solutions near the neutral point are plotted in Fig. 1. The adsorption on either mineral was far greater for the aminoethyl starch than for the Pearl starch. Flocculation The degree of flocculation was determined for a given pulp by measuring the time required for the supernatant liquid to reach a certain degree of clarity. This was accomplished by passing a standard light beam through the pulp, contained in a glass settling tube, and measuring the time required after agitation ceased for the intensity of the transmitted beam to reach a predetermined value. Starch reagent was added to each pulp in small uniform increments, and the settling time was determined after each such addition. Increase in pulp liquor volume due to the added starch reagent was ignored in these tests. Pulps were maintained at or near neutrality. of the preparation of the aminoethyl starch: "Three hundred and sixty grams of powdered unmodified corn starch was slurried in 3 liters of water and stirred during gelatinization. Eighty-six grams of ethylenimine was added slowly during reflux which was continued for 4 hr. The yellow solution set to a gel upon cooling. This was broken up and neutralized with HC1. The thick paste was dried on hot rolls. Adsorbed polyethylenimine and salts were removed from the yellowish products by washing first with acidified aqueous '60 per cent' alcohol. The white product contained nitrogen corresponding to about one aminoethoxy group per ten glucose residues." Only laboratory samples of this starch derivative have been prepared. The starch reagents for use in flocculation, flotation, and adsorption tests were made up by dispersing the appropriate quantities of air-dry starch in cold water, adding boiling water, and heating for 30 min at 120°C under pressure. When used ip adsorption tests, the starch solutions first were cleared of insoluble matter by centrifugation.

Jan 1, 1953

By Fred C. Bond

DISCUSSION Milton C. Shaw and Donald R. Walker (Massachusetts Institute of Technology)—For many years the two empirical rules of comminution bearing the names of Rittinger and Kick have been applied to grinding processes to account for the energy required to crush brittle materials. The basis for Rittinger's rule was the thought that the energy to grind a substance went largely into supplying the additional energy associated with the new free surface, while Kick's law was prompted by the concept of dimensional similarity and is a special statement of the concept of constancy of strain energy to cause rupture. According to Rittinger the energy to grind a given volume varies as the change in the square of the mean particle size, v - A (D2), while according to Kick this energy varies

Jan 1, 1953