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By Charles A. Colton

(Read at the Amenia Meeting, October, 1877.) THE results of a series of analyses extending over a period of three weeks at the Cedar Point Iron Company's furnace, Port Henry, New York, are given in Tables I and II. This furnace uses a very pure magnetite and Lehigh and Lackawanna anthracite. The flue leading the gas from the "down-comer" to the boilers was tapped by a a inch gas pipe, which carried the gas to the Orsat apparatus. The pressure of the gas not always being sufficient, owing to the small amount made, to force it into the apparatus, I omitted taking samples several days, and with the exception of the last three days of the campaign, made the analyses whenever opportunity offered. I find the Orsat apparatus, as described by Prof. Egleston,* to work very well, with one exception. The CO is not absorbed as readily in the ammonia-copper solution as he states in his description of the apparatus, as many as 50, and sometimes 60, passes being necessary to absorb all the CO. The power of absorbing rapidly increases as the solution is used, and this would indicate that the more oxygen it contains the quicker will it do the work. Probably the addition of a small amount of some oxidizing agent would remedy this. When the furnace was working in its normal condition, I had no occasion for using the filter, the amount of fume being so small as not to cause any inconvenience. The gas burned with the flame peculiar to CO, and contained just enough solid particles to give it a slight reddish tinge. * See Transactions, II, 226; V, 487, 621.

Jan 1, 1878

By David R. Kalasky

One of the primary areas of computer application in mine ventilation has been the analysis of mine ventilation networks. A variety of network analyzers that are aimed toward the solution of flows in a given network has been developed. These programs typically use the Hardy Cross or other iterative algorithm for solution of any network, even those that are complex. The computer software has become more efficient and sophisticated over time and has thus been widely adopted for ventilation network analysis. One of the shortcomings of these programs is that they have been developed with the idea of analyzing the ventilation requirements at a single point in the lifetime of a mine. This provides the ventilation engineer with an important tool for analysis but falls somewhat short of providing a convenient means to economically com- pare alternative methods for ventilating a mine. The program described in this paper provides such a tool. The principal concepts that were necessary for the completion of a computer program that analyzes ventilation alternatives is the capability of efficiently determining the ventilation requirements at a variety of times during the lifetime of the ventilation system and some method of analyzing the costs of the alternatives on a realistic comparative basis. To accomplish this task, the authors began using Penn State's mine ventilation simulator designated as PSU/MVS, Didyk et al. (1977a, 1977b) and added the necessary additional software to accommodate the need for multi-period, time-dependent analysis of the net- work. The software was then tested on a coal mine ventilation problem with a 23-year life. The problem points out the value of such a program for ventilation analysis.

Jan 1, 1986

By Junji Shibata

As one of the effective usages of various wastes discharged in non-ferrous metal industry, the synthesis of hydrotalcite was carried out using aluminum dross as a raw material. Hydrotalcite is one of the inorganic anion exchangers and has layered structure of complex hydroxide. The removal of toxic metal ions such as As(III), As(V), Cr(VI) and Se(IV) with the hydrotalcite was investigated, and the removal properties were compared from the viewpoint of the anion exchange amount and exchange rate. Hydrotalcite can be obtained by a co-precipitation method from the aluminum dross and the waste MgCl2 solution discharged in an aluminum regeneration process. Slight amounts of SiO2, Al2O3, Fe2O3 and so on remain in the reaction product as an impurity component. Interlayer distance of the obtained hydrotalcite is about 0.3nm. Hydrotalcite changes to Mg-Al oxide by the calcination at 773K for 3h, and then hydrotalcite is formed again by a rehydration operation after calcination. The toxic metal ions are removed from aqueous solution by the anion exchange reaction in the pH region where toxic metal ions exist as the anionic species like AsO2-, HAsO42-,CrO42-, and SeO32-. The exchange amount is considerably concerned with the valence of anionic species and the size of anionic species. The removal of toxic metal ions were investigated from the engineering aspects such as the low concentration limit to be removed, the amount of hydrotalcite to be used for the perfect removal of toxic metal ions and so on.

Jan 1, 2006

By C. H. Herty

The reactions that occur in the basic open-hearth process are complicated and are subject to many different factors such as temperature, slag composition and rate of boiling of the bath. It is difficult, therefore, to present an extended theoretical discussion of open-hearth operations without confusing the theory of the process with the application of the theory. This paper attempts to present some of the theory of the open-hearth process without mentioning theory at any point, simply giving some of the practical applications of results founded both on theory and practice and on experimental work both in small and large furnaces. It is hoped that an answer to a number of questions commonly asked by the open-hearth operator may be found in this paper. Discussion of the subject matter is taken up more or less in the order in which a heat ie worked, beginning with the oxidation of the scrap charged, sulfur absorption by the scrap, calcination of the limestone, and extending into the pig iron for the charge, the slags formed during the process, and some of the phenomena occurring in deoxidation of the steel and in its action in the molds. Oxidation of Scbap When scrap is charged into a furnace it immediately begins to oxidize, Oxidation is carried out by.three reactions: Fe + H20 + FeO + Hz (absorbs heat) (1) Fe + C02 + FeO + CO (absorbs heat) (2: Fe+ ½O2 + FeO (gives off heat) (3: The product of oxidation contains some Fe3O4, but FeO usually predominates on account of the iron present in close contact with the scale formed For a given type of scrap the amount of oxidation is controlled by thc fuel and by the excess air being carried in the furnace, the temperature 01 the furnace, and the length of time that the scrap is in contact with the furnace gases. In general, less oxidation occurs when the furnace is rur hot so that the scrap melts rapidly. More important in operation thar the two variables atmosphere and temperature, is the type of scrap being

Jan 1, 1929

WE vote this month without knowledge of either major presidential candidates' position on mineral problems. Too long have we kept our own counsel, turned our backs on the public. Certainly we must formulate policies upon which we may unite in industry council, but implementation of these by legislative bodies requires popular understanding, if not support. The earth and its resources are the province of the miner and the farmer but beyond this, and the basic need of the economy for both, the similarity ends. In this election year the disparity between the poll strength of these two fundamental callings is apparent. While both candidates have found it essential to incorporate a farm plank in their platforms, mining has been neglected as not being a basic issue. If this gives the mining fraternity the advantage of voting objectively, it also leaves the general public in the dark as to the attitude of the candidate on a basic issue. It is an issue on which depends the future of the nation, but, unfortunately, the public is not aware of its own vital stake in mining. Thus election time brings with it a fresh example of the evils of automatic lawn sprinkling-a long standing peeve of ours. Back in the days of the hand-held sprinkling hose in the intimacy of the twilight the mining man had the opportunity to unburden himself on the problems of mining to his neighbor who might be a doctor, lawyer, or accountant. With the encroachment of mechanized sprinkling, yet another forum for the indoctrination of the layman into the significance of mining has disappeared. Familiarizing your neighbors with mining in the lawn sprinkling societies of yesteryear was the individual's contribution to public relations for mining. The corporations can use educational advertising in popular magazines as is practiced by such companies as Kennecott. The AIME can play an important part by providing career guidance information on the mineral industries. We offer the suggestion to the American Min¬ing Congress, that the splendid exhibitions of mining machinery featured in alternate years for metal and coal mining, respectively, might be opened to the general public one night during the week of the show. Newspaper publicity would add greatly to the public's interest in such an opportunity. Power for the public is considered a vote-getter and no candidate has neglected to make his stand on this subject. Is the price of the kilowatt in the home of greater importance to the voter than the price of the metal durable goods which he purchases every year? Not really, but the voter thinks so. Mining has as much sales appeal as 'potatoes or kilowatts if presented in the proper way. We suggest that a little hand lawn sprinkling by the right people in the right places would produce an abundance of publicity which would give Mr. and Mrs. American something new to expect from their representatives-a vigilance over the flowering of their mining industry.

Jan 1, 1952

Study on Iron Ore Sintering with High Proportion of Specularite Concentrates Specularite iron ores have poor hydrophilicity and ballability as well as bad high-temperature reactivity, which restricts the large-scale application in the sinter and pellet production. In order to make the best use of the specularite concentrates, a novel process of Composite Agglomeration Process (CAP) is applied in this research. By the CAP, the percentage of specularite concentrates in the sinter mixture is increased to no less than 40 per cent as well as the coke breeze dosage is reduced by 0.3 per cent~0.6 per cent. Moreover, all the sintering indexes are obviously improved compared with the traditional sintering process. Researches on the metallurgical performances shows the finished composite agglomerates are high-quality burdens for blast furnaces. This investigation can provide an effective route to utilize the inferior specularite concentrates in the sintering.

Sep 13, 2010

By David Cole

THE appearance of the Wilfley table in 1897 marked an epoch in the art of concentration of ores. The table has merited and received an almost unprecedented measure of public approval, lasting through its whole patent life. It has been very little improved in itself, or improved upon by competitors. The new machines bidding for popular favor have been of the Wilfley family, but without ability to do markedly better work than the parent m:-chine, or to (to more work except by the use of superimposed decks, the latter having obvious disadvantages. FIG. 1.-The- Butchart RIFFLE System. During the past three years there has been developed at Morenci a new type or arrangement of riffles applicable to the Wilfley type of concentrating table which corrects many of the objectionable features or limitations of the older system and obviates most of the difficulties encountered with it, particularly when handling ores having a low ratio of concentration. The new type is known as the Butchart1 riffle system, and its general arrangement is shown in Figs. 1 and 2.

Jan 2, 1915

By J. Frenzel

The computation of hydrometallurgical equilibria is very difficult because of the many complex species in solution, the precipitates, condensed species, or solids, and the activity coefficient corrections necessary to non-ideal thermodynamics. Our objective was to tryout recently developed genetic algorithms to minimize free energy and solve the equilibrium computation. We selected test problems that were seemingly unsolvable with more traditional algorithms, and we applied genetic algorithms to them. We were pleased to find that genetic algorithms succeeded where other algorithms did not. We concluded that they are an effective tool for hydrometallurgical equilibria.

Jan 1, 1992

By David Keys

THE object of this investigation was to determine with a vertical magnetic variometer the extent of the titaniferous ore deposit that occurs not far from Ivry in Terrebonne County, Quebec, about 70 miles north of Montreal. The Ivry mine, on lots 37 and 38 of range V, Beres-ford township, Quebec, is one of the largest and most accessible to transportation of the deposits of ilmenite in Quebec. The ore consists of an aggregate of ilmenite anhedra serving as host for numerous fine plaques of hematite. No magnetite is known from any occurrence near this lot, and in this way the deposit differs from many others that contain a large percentage of magnetite with the ilmenite. The deposit at Ivry thus has much lower magnetic susceptibility than those containing abun-dant magnetite, and a sensitive magnetometer was necessary for the work, although a body of magnetite with ilmenite about a mile from the Ivry mine was successfully prospected with a dip needle. The principal orebody is exposed in a quarry known as the Ivry mine, from which about 16,000 tons of ore have been shipped. Small exposures of ilmenite isolated by drift serve to show the presence at the surface of some of the bodies indicated by the magnetometer,- but most of the hill is covered with glacial till and second growth hardwood. At the lower levels the till gives place to fluvioglacial sands and gravels from which the forest has been cleared. The increasing thickness of the sand and gravel as the hill descends makes the tracing outward of the deposits from the rocky knob of anorthosite a difficult matter. Within the orebodies too, some masses of anorthosite persist, which may account for some local irregularities. Fig. 1 is a photomicrograph, taken by reflected light, of a typical specimen of the ore. The dark patches are hematite and the light parts are ilmenite. No trace of magnetite was found. Fig. 2 is a photo-micrograph of a thin section of the country rock showing the texture of the anorthosite. The mean of four analyses quoted by Robinson1 is: titanium, 19.26 per cent; iron 45:41 per cent.

Jan 1, 1936

R. H. Sweetser.—Mr. Reinartz spoke of slag containing 18 per cent FeO. I think he expressed the iron as oxide. I should like to ask at what point, in iron contents, it is usual practice to quit sending that open-hearth slag to the blast furnace? L. F. Reinartz.—NO problem is involved in disposing of open-hearth slag to our blast furnaces because, on account of the grades of iron and steel made by our company, the iron oxide in the slag is sufficiently high to make it very acceptable to the blast-furnace operator. I have been told that if the total iron oxide in the slag is under 18 to 20 per cent, it is not economical to use such slag in a blast furnace. J. T. Mackenzie,* Birmingham, A1a.—Mr. Reinartz mentioned sulfate in the limestone. I wonder if he did not use the wrong word. We always have assumed that the sulfate sulfur is practically harmless. Most limestone carries some pyrite (sulfide sulfur) which we thought was what gave the sulfur pick-up. L. F. Reinartz.—I have had no experience, personally, in the use of limestone high in sulfates. I have been told, on several occasions, by other open-hearth men, that the sulfates will be reduced and increase the sulfur content of the metal. We have never found any pyrite in our limestone. A. Hayes,† Middletown, Ohio.—I should think that sulfates would be readily reducible to sulfites, and possibly to sulfides at various distances through the slag layer, from the top down to the interface between the bath and the slag, because reducing conditions are certainly sufficiently strong next to the bath in the slag to reduce sulfates to sulfides. A. L. Feild,‡ Baltimore, Md.—What is Mr. Reinartz's theory as to the effect of silica in the slag on the iron oxide content of the metal? L. F. Reinartz.—I have no particular theory as to the effect of silica in the slag on the iron oxide content of the metal. I know from practical experience that it does have an influence on reduction of iron oxide in the metal. We know that if the silicon content of the charge, and, therefore, the silica of the slag is increased, the iron oxide in the slag will be decreased. Based on Dr. Herty's work, we assume that the iron oxide is correspondingly decreased in the metal. C. H. Herty, JR.,§ Bethlehem, Pa.—In the theory of iron oxide control, we have tackled the matter from the physical standpoint, rather than the chemical standpoint or from the slag itself, and we worked on the basis that the change in iron oxide content in a slag—that is, whether the slag increases or decreases in iron oxide—is due to the simple fact that if you put more iron oxide into the slag from the gas than you put oxygen from the slag into the metal, the slag must increase in iron oxide;

Jan 1, 1935

R. H. Sweetser.—Mr. Reinartz spoke of slag containing 18 per cent FeO. I think he expressed the iron as oxide. I should like to ask at what point, in iron contents, it is usual practice to quit sending that open-hearth slag to the blast furnace? L. F. Reinartz.—NO problem is involved in disposing of open-hearth slag to our blast furnaces because, on account of the grades of iron and steel made by our company, the iron oxide in the slag is sufficiently high to make it very acceptable to the blast-furnace operator. I have been told that if the total iron oxide in the slag is under 18 to 20 per cent, it is not economical to use such slag in a blast furnace. J. T. Mackenzie,* Birmingham, A1a.—Mr. Reinartz mentioned sulfate in the limestone. I wonder if he did not use the wrong word. We always have assumed that the sulfate sulfur is practically harmless. Most limestone carries some pyrite (sulfide sulfur) which we thought was what gave the sulfur pick-up. L. F. Reinartz.—I have had no experience, personally, in the use of limestone high in sulfates. I have been told, on several occasions, by other open-hearth men, that the sulfates will be reduced and increase the sulfur content of the metal. We have never found any pyrite in our limestone. A. Hayes,† Middletown, Ohio.—I should think that sulfates would be readily reducible to sulfites, and possibly to sulfides at various distances through the slag layer, from the top down to the interface between the bath and the slag, because reducing conditions are certainly sufficiently strong next to the bath in the slag to reduce sulfates to sulfides. A. L. Feild,‡ Baltimore, Md.—What is Mr. Reinartz's theory as to the effect of silica in the slag on the iron oxide content of the metal? L. F. Reinartz.—I have no particular theory as to the effect of silica in the slag on the iron oxide content of the metal. I know from practical experience that it does have an influence on reduction of iron oxide in the metal. We know that if the silicon content of the charge, and, therefore, the silica of the slag is increased, the iron oxide in the slag will be decreased. Based on Dr. Herty's work, we assume that the iron oxide is correspondingly decreased in the metal. C. H. Herty, JR.,§ Bethlehem, Pa.—In the theory of iron oxide control, we have tackled the matter from the physical standpoint, rather than the chemical standpoint or from the slag itself, and we worked on the basis that the change in iron oxide content in a slag—that is, whether the slag increases or decreases in iron oxide—is due to the simple fact that if you put more iron oxide into the slag from the gas than you put oxygen from the slag into the metal, the slag must increase in iron oxide;

Jan 1, 1935

By W. H. Schacht

The operations of the Copper Range Co. are located in the Michigan copper district at the southern end of the Keweenaw Peninsula, 8 miles southwest of Hough ton (the center of the district). All the producing mines are situated in a long narrow zone paralleling the peninsula and occupying the central portion. This zone is about 25 miles long and 3 to 4 miles wide. The important mines, in their order from northeast to southwest, are, Mohawk, Ahmeek, Allouez, North Kearsarge, Wolverine, South Kearsarge, Centennial, Calumet & Hecla, Osceola, Franklin, Quincy, Isle Royale, Baltic, Trimountain, and Champion; the last three are operated by the Copper Range Co. The existence of copper deposits in this district became known soon after the discovery of America. Many references by early explorers show that copper was found in the possession of many of the American Indians that inhabited the Great Lakes region. There is abundant evidence that exploration of these copper deposits extended into prehistoric times; and some indication that the American Indians carried on operations shortly before the advent of white men. An account of the finding of primitive mining appears in the Government publications of 1850-51 by Foster and Whitney. Although earlier expeditions by white men had been made to the Keweenaw peninsula, an attempt to commercialize the district was not made until after 1765, when Alex. Henry, a British subject, searched the shores of Lake Superior for mineral treasure. His first expedition took him to the Ontonagon River, where he found the famous mass of copper now on exhibition at Washington. In 1771, he started operations along this river with a crew financed by English capital; but after much difficulty and only meager success he found only detached masses while in search for a major deposit. There is no record of any cxploration from 1774 to 1820. In 1838, Dr. Douglas Houghton was appointed state geologist. Prior to that time he had made several visits to this region; and was the first white man to find

Jan 1, 1925

By C. J. Hartley

"Scandium oxide can be recovered as a by-product from process streams in a number of chemical and metallurgical plants including TiO2 sulphate pigment, Ni/Co laterite, zirconium oxychloride and uranium operations. The feed materials available for processing can be either mid or end of process streams containing many species in solution. This paper describes the strategies available to limit coextraction of elements into scandium containing streams from TiO2 pigment production and the removal of these elements to deliver high purity scandium oxide.A critical part of process development is to work within the constraints of the main process in terms of introducing new unit processes and producing minimal additional process intermediates and residues. This requires flexibility in designing flowsheets and plants to find novel process routes that avoid production of intermediates and/or the introduction of reagents that can compromise the main process operations.Bloom Energy has developed novel solvent extraction processes for the extraction and refining of scandium oxide that minimise impacts on core process streams within TiO2 pigment plants. The additional unit processes are designed to have minimal contact with pigment plant process streams to produce a concentrate that is refined to nominal 3N purity scandium oxide in a simple, small scale commercial plant."

Jan 1, 2016

By V. Yahorava

Efficient recovery of precious metals from process solutions is essential for improving process economics. Traditionally, precious metals are relatively effectively recovered from waste streams via precipitation or cementation. However, these approaches have a number of drawbacks, including poor water balance, creation of environmentally unfriendly waste streams, and losses of precious metals. Ion exchange technology is an alternative for the recovery of precious metals from waste or low-grade streams. This technology allows the recovery of the precious metals to extremely low levels (micrograms per litre) with relatively high upgrade ratios from the solution onto the resin without major water balance concerns, while the impact on the environment could be minimized or avoided. Research was conducted on the recovery of platinum group metals and gold from different low-grade and waste streams from one of the precious metals refineries in South Africa by means of ion exchange. Various functionalities and matrices (granular and fibrous) of ion exchange materials were evaluated. The results from these studies indicated that in some cases ion exchange could be very effective for the recovery of precious metals, and that the PGM concentration could be reduced to < 1 mg/L. The upgrading ratios of the various PGMs onto the specific fibres were relatively high for the specific streams evaluated, which might in some cases justify incineration of the loaded material instead of stripping and recycling the adsorbent. The cost of direct incineration for one of the waste streams tested would be less than 1% of the value of the PGMs recovered. However, the adsorbent has to be carefully selected and the process design optimized for each specific stream.

Jan 1, 2014

AS you and anyone else can understand, a great expense and a large supply of things are necessary for parting a quantity of silver by means of aqua fortis. First, as you have seen, it is necessary to have a great abundance of every kind of cucurbit and alembic, of charcoal and wood, and a large number of pounds of a vigorous, well-purified aqua fortis, because for every pound of silver there are of necessity [71v] at least four pounds or more employed from the parting to the coloring of the separated gold. It is also necessary to prove its fineness and to granulate or beat it, to have furnaces and supplies, and to provide for a thousand other necessities which need not be repeated now. The most important thing in this art seems to me to be that it requires continual working night and day with extreme watchfulness and care. For this reason another way, much quicker, less dangerous, and less expensive than the aforesaid one was found by some clever men (alchemists, I believe) in order to avoid such laborious and heavy methods. It produces the substance of the thing exactly as the acid does. If you wish to make use of it in conjunction with the acid it is possible to render great aid to this work although there is a great difference in the techniques of the two. The method used in this parting is as follows: First, make a round or square wind furnace for melting, below ground or above, and of height and size in proportion to the work and as it comes out best for you. Then take a large clay crucible and fill it almost up to the edge with that silver which contains the gold, cut into pieces, and put it to melt with the silver on top of a piece of brick cut to the size of the bottom of the crucible and resting on a grate in the middle of coals that are burning well. When you see that the silver has been made so hot by the fire that it looks white and that it wishes to begin to liquefy, take a little stick of sulphur or a small piece of antimony and put it in. When this is melted, lifting away the charcoal that covers it, put in more sulphur or antimony (for the method is the same with either one) until the silver is very well melted and these materials properly incorporated. Then for every pound of silver that you put in add half an ounce of sheet copper. When this is melted take your crucible out with tongs or pincers and strike the bottom gently on the coals two or three times so that the gold may make a residue on account of

Jan 1, 1942

By A. Kenneth Schellinger, O. Cutler Shepard

INTRODUCTION OVERGRINDING, or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncontrolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about 90 pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, overgrinding is detrimental to the flotation process. Overground slime coats larger particles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 Conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The remedy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. DEVELOPMENT OF APPARATUS AND EXPERIMENTAL METHOD Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7 ½ in. in diameter and 8 in. deep. A cast-iron disk was mounted on the end of a vertical shaft so that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffeemill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, weighed charges of ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned on so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1947

By A. Kenneth Schellinger, O. Cutler Shepard

Overgrinding, Or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncon rolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about go pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, over-grinding is detrimental to the flotation process. Overground slime coats larger partitles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The rem- edy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. Development oF Apparatus and Experimental Method Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7½ in. in diameter and 8 in- deep. A cast-iron disk Was mounted on the end of a vertical shaft SO that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffee-mill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, more charges of Ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned On so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1949

By J. E. Owen

EXPERIMENTS on twenty-six drilling muds were carried out in an effort to establish a correlation between the apparent viscosity values for these muds as obtained with the Marsh funnel and the Stormer viscometer. This required a calibration of both instruments in which the density of the calibrating liquids was taken into account. Corresponding to the Marsh-funnel viscosity values that were obtained at 1500 c.c. in, 1 qt. out, equivalent Stormer viscosity values were obtained at a Stormer cylinder test speed of approximately 1100 r.p.m. Further, a linear relation between the logarithms of the Stormer cylinder test speed and the resulting apparent viscosity was found to exist for the muds tested for 200 to 1000 r.p.m. The significance of the slope of this linear relationship is discussed as it applies to the field use of drilling muds. It was concluded that Marsh-funnel viscosity measurements are expressible in absolute units that may be correlated with Stormer viscosity measurements, provided these are made in the manner described in the paper. INTRODUCTION The Marsh funnel is used almost universally to determine the viscosity of drilling muds in the field. The Stormer viscometer is commonly used to determine the viscosity of drilling muds in the laboratory. With the Marsh funnel, the time required for one quart of mud to flow out of the funnel originally filled with 1500 c.c. is often taken as the measurement of viscosity. With the Stormer instrument, drilling mud is tested at a cylinder speed of 600 r.p.m. and by means of calibration data the apparent viscosity is expressed in absolute units. Neither of these instruments is an absolute viscometer, and apparent viscosities obtained with either are influenced by the arbitrarily standardized conditions under which the measurements are made. It is disconcerting to the man in the field to be told that laboratory measurements of the viscosities of a group of muds differ widely, when he has already observed that they have identical viscosities when tested in the Marsh funnel. Drilling muds are not perfect fluids, therefore it is not surprising to find a lack of correlation between the apparent viscosity of a mud as determined with the Marsh funnel and that determined with the Stormer viscometer. The work described here was undertaken to develop a correlation between Marsh-funnel and Stormer viscosity measurements. To a considerable degree this has been accomplished. This study has resulted not only in a better understanding of the two viscosity instruments and the data obtainable from them, but has given considerable information regarding the viscosity characteristics of drilling muds. EXPERIMENTAL PROCEDURE Calibration of Harsh Funnel and Stormer Viscometer The first step in the experimental procedure was the calibration of the Marsh funnel and the Stormer viscometer with

Jan 1, 1941

By CHARLES L. WRIGHT

Coals may be divided into six classes-anthracite, semi anthracite, semibituminous, bituminous, subbituminous, and lignite. The first three cla~ can be distinguished by dHferences of composition, particularly the proportions of fixed carbon and the ratios of the fixed carbon to the volatile matter in the cools. The last three classes can be distinguished by di:fferences in physical character, chiefly in color and in manner of weathering. Lignite is brown, not black, and has generally a woody look, but it weathers in much the same way as subbituminous coals. On exposure to the air lignite slacks or crumbles. The lumps check and fall into small irregular pieces that exhibit a decided tendency to separate into extremely thin plates. Hence lignite deteriorates greatly during storage or long transportation. The most characteristic feature of the composition of lignite is a large percentage of moisture. '!his high moisture content reduces the fuel value of freshly mined lignite, and the partial evaporation of moisture, on expooure, causes th6 fuel to check and fall to pieces. Consequently, attempts to increase the efficiency of lignite as a fuel involve reducing its percentage of moisture and increasing its ability to endure storage and transportation. Both these results are accomplished by briquetting.

Jan 1, 1911

By J. J. Taute, O. S. Tinkler, G. Zárate, K. C. Sole, J. Steeples

Global practice in copper solvent extraction is reviewed, based on individual plant operating data for 2018. The survey respondents account for almost 40% of global copper cathode production by solvent extraction and electrowinning. Regional trends in operating conditions, as related to the prevailing ore mineralogy, leaching processes, and life of mine are evaluated. The adoption and expansion of this technology in the African Copperbelt and Asia are assessed, together with the associated challenges of specific orebody characteristics. Trends in feed and electrolyte compositions and flow rates, and extractant and diluent selection, together with their impact on copper recovery are reviewed. The evolution of equipment design and associated ancillary processes are discussed, and new projects identified. INTRODUCTION Continuing from similar surveys carried out in 1997, 1999, 2003, 2007, and 2013 (Robinson, Davenport, & Jenkins, 1997; Jenkins, Davenport, Kennedy, & Robinson, 1999; Robinson, Sandoval, & Cook, 2003; Robinson, Moats, Davenport, Karcas, Demetrio, & Domic, 2007; Sole, Zarate, Steeples, Tinkler, & Robinson, 2013), global operating practice in copper solvent extraction (SX) in 2018 is reviewed. Contributing operations include established sites in the USA (8), Chile (9), Peru (2), Spain (1), Mexico (2), and Australia (2), as well as newer operations in Africa (7) and Asia (3). This survey represents 2018 copper cathode production of 1.91 Mt, or about 43% of world production by solvent extraction combined with electrowinning (SX–EW). Trends in operating conditions and performance, as well as equipment design and process technology choices are analysed. The significant expansion in production in the African Copper Belt is also discussed. There are currently about 130 operations worldwide producing some 4.6 Mt copper by SX–EW; of these, about half produced more than 10 000 t/a. Figure 1 shows global copper production by country. As it has for several decades, Chile remains dominant, with 36% of global production of cathode copper. Since 2005, the Democratic Republic of Congo (DRC) has increased production year on year, and now represents 23% of the global total (International Copper Study Group, 2017; Statistica, 2018). Copper production in Asia has grown significantly in the past decade, particularly since the ramp-up of Wanbao in Myanmar. There used to be numerous small copper SX plants in China (Schlesinger, King, Sole, & Davenport, 2011), although most of these have now shut down and Chinese production is lower than it was five years ago. The most important trend, however, is the rise of Chinese-owned mining companies, particularly in Central Africa, but also around the world more generally.

Jan 1, 2019