Search Documents

By R. W. Smith

It is well known that certain silicates such as the feldspars and beryl can be readily floated from quartz at a low pH value using a long chain cationic collector in the presence of a fluoride.1-3 However, the mechanism whereby the fluoride activates feldspars and beryl is not known. Indeed, only a few authors have even speculated about possible mechanisms.1,3,4 As a first approach to the problem, it is perhaps best to look closely at the interesting character of hydrofluoric acid and how it dissolves silica and silicic acid. Hydrofluoric acid, unlike hydrochloric acid, is a weak acid with a dissociation constant of about 3.53 x 10-4 at 25°C.5 The reason for the weak acidic character of HF lies in the very small size of the fluoride ion.6 The pKa value is 3.45 which means that at pH 3.45 half the HF is present in molecular form and half in ionic form. The anion present is probably HF2.-6 At pH values below the pKa value, molecular concentration increases rapidly at the expense of the HF2- anion. It is known that HF readily attacks glass, silicates, quartz and silicic acid. The equation usually cited pertaining to the first step in HF attack of silica is6-'

Jan 1, 1965

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

By H. Rush Spedden, Arvid Thunaes

R. R. Knobler and F. E. Albertson—Following the testwork done by Thunaes and Spedden, a Sullivan deck plant was built for the Colquiri mill. This plant started to operate in April 1945 and continues in successful operation. Since May 1950 the Colquiri mill has operated with a completely revamped flowsheet at a rate of 1000 tons per day. The old Sullivan deck section was retained with this new flowsheet. The "slime plant" in Colquiri consists of one 12-cell Denver No. 18-Sp. Sub-A flotation machine, one 16-ft and one 12-ft hydroseparator, six five-deck Sullivan-frames and six shaking tables for cleaning up the Sullivan deck concentrates, as well as 12 Deister plus three Plat-O tables on which the fine sands, which are not treated on the S-decks, are concentrated. Only —325-mesh fines are treated on the Sullivan decks. The combined hydroseparator overflows are fed to the S-decks. From 1946 to 1949, i.e. with the "old" flowsheet and an average mill feed of 800 tons per day an average of 45 tons per day was treated on four Sullivan rougher frames. The two additional frames were used as cleaners. At times, up to 90 and 120 tons of slimes were treated successfully on this Sullivan deck installation. The average operating data for the S-deck section in Colquiri for the year 1949 were as follows: The tests reported in this paper indicated a tin recovery of 42 pct. The actual plant results for 1949 show a 54 pct tin recovery from the Sullivan decks. The total average quantity of tin produced from the S-deck section in 1949 was 15 tons of tin per month. This corresponds to 3 pct of the total mill recovery. The difference between the 6 pct overall recovery-increase expected after the tests in 1943 and the actual plant results given for 1949 is explained as follows. According to the tests, a total of 170 tons was expected to be routed to the Sullivan decks, but because of certain changes in the flowsheet and mill amplifications,

Jan 1, 1951

By R. S. Olson

An analysis of the wash thickener-two stage filtration circuit was undertaken to quantitatively describe the circuit performance. On the basis of information developed in this study, the factors of major importance in reducing soluble loss are: filter wash, wash zone size, filter capacity, filter feed pulp dilution, blow back, filter cloth, water to repulp, water to primary wash, water to feed pulp and thickener underflow. The wash thickener-two stage filtration circuit has been used in many hydrometallurgical plants to separate solute from leach solids. As many as 20 to 30 plant operating variables determine the performance of the wash circuit and hence the soluble loss. The complex relationships of these variables, with each other and with soluble loss, make the evaluation of their effect on mill performance difficult. An analysis of this type of circuit was undertaken to quantitatively describe the circuit performance. THE CIRCUIT The diagram in Fig. 1 represents the operation of the wash thickener-two stage filter circuit. This circuit provides for counter current, cross current or for a combination of countercurrent-cross current operation. The feed pulp stream, J, is intended to include any solution or water, other than primary and secondary filtrate, which enters the wash thickener. Except for filter washing, the operations in the circuit are simple mixing operations which can be described by material balance. The concentration of solute in the filtrate and cake solution after application of wash liquor cannot be described by simple material balance. The displacement washing equation, derived by Rhodes1 and extended by Choundhury and Dahlstrom,2 appears to best describe this case of miscible displacement in porous media. In a useful form, it is

Jan 1, 1963

By Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

0. C. Ralston and F. Fraas—Dr. Sun and associates have presented an interesting paper not all of which is comprehended by us. The data assembled measure the deflections of particles in an electrostatic field as a function of a number of independent variables and some dependent variables that are not sharply differentiated. These data are all based on a Johnson type machine of definite, well-described geometry, something not often done in electrostatic separation literature. One new fact brought out by this technique is the effect of coal dust on admixed pyrite and quartz. The effects are opposite in character, as should be expected and we do not agree with the authors that these effects are negligible. Fraas' also used a multiple cell "distribution analyzer" and gives in fig. 5 of his paper a straight line plot with no humps or curves. This is not necessarily at variance with Sun's results because Fraas used a larger gap between electrodes and had no evidence of particles adhering to or dropping off the charged electrode. The section of Sun's paper on effect of surface conductivity contains a speculation that the dielectric constant "represents more or less the electrical conductance of the bulk body instead of the surface of the mineral particles." A simple picture of the meaning of the dielectric constant is that it is the specific inductive capacity of a dielectric when used as the dielectric between the plates of a condenser. It is at once evident that the above speculation confuses capacity with conductance—two definitely independent variables. We ask the authors to state in what group or subdivision their garnet belongs; what method and units were used in calculating the data of col. A, table I and their meaning; what was the temperature of the carrier roll and, finally, has any effort been made to investigate the effects of particle shape on distribution in the electrostatic field? S. B. Hudson—I have read this article with great interest. We have been engaged in research work on the principles of electrostatic separation in this laboratory for some time now, and our findings agree with those of the authors in many respects. The article shows evidence of careful and valuable research in the field of electrostatic separation. A "distribution analyser," very similar to that described in an earlier article by one of the authors," was incorporated in an inclined plate-type electrostatic separator designed and built in the Melbourne University laboratory in 1948 for investigation purposes.22 The actual splitting edges were machined from y! in. per-spex, and the paper hoppers were supported on linen thread immediately below the perspex dividers. These dividers fitted into machined slots in a framework to give accurate Yz-in. spacings. The hoppers (staggered) fed directly into a rack of test tubes, which is supported on a vertical pantograph arrangement. The rack was positioned with guides on the horizontal pantograph stand, and this ensured positive alignment when replacing the rack after making weighings. In later work, when much heavier feed rates were used, of the order of 30 to 40 Ib per in. per hr a rack fitted with rectangular metal containers and similarly aligned was used. Some work was done here on comparing the distributions of minerals when passed separately and when passed as a mixture, and it was found that there was quite an appreciable difference in the two results.= However, in our separator the particles do not pass down the plate in a single layer, and this difference is probably caused by collisions of one mineral particles with the other mineral particles. In most of the investigational work here, the change of the center point of the distribution is measured to establish the effect of a variable, such as voltage. Two minerals (zircon and rutile) have been studied rather exhaustively, and it was found that their distributions are very nearly normal. Owing to the sharpness of the distribution curves, the usual method of obtaining the mean or median was inaccurate, and was not used; instead the mean (also the median), calculated on the assumption of a normal distribution, was used to locate the center point of each distribution and proved satisfactory. The effect of polarity becomes very apparent in the plate-type separator where frictional charges play a very important part when using highly resistive minerals such as zircon. With rutile, a comparatively conductive mineral, polarity of the electrode has little effect. On the other hand, the magnitude of the voltage has a far greater effect on conductive than on resistive minerals. Shiou-Chuan Sun (authors' reply)—Thanks are extended to Drs. Ralston and Fraas for their keen interest in this paper. Their questions concerning coal dust,

Jan 1, 1951

By Shiou-Chuan Sun, R. F. Wesner, J. D. Morgan

0. C. Ralston and F. Fraas—Dr. Sun and associates have presented an interesting paper not all of which is comprehended by us. The data assembled measure the deflections of particles in an electrostatic field as a function of a number of independent variables and some dependent variables that are not sharply differentiated. These data are all based on a Johnson type machine of definite, well-described geometry, something not often done in electrostatic separation literature. One new fact brought out by this technique is the effect of coal dust on admixed pyrite and quartz. The effects are opposite in character, as should be expected and we do not agree with the authors that these effects are negligible. Fraas' also used a multiple cell "distribution analyzer" and gives in fig. 5 of his paper a straight line plot with no humps or curves. This is not necessarily at variance with Sun's results because Fraas used a larger gap between electrodes and had no evidence of particles adhering to or dropping off the charged electrode. The section of Sun's paper on effect of surface conductivity contains a speculation that the dielectric constant "represents more or less the electrical conductance of the bulk body instead of the surface of the mineral particles." A simple picture of the meaning of the dielectric constant is that it is the specific inductive capacity of a dielectric when used as the dielectric between the plates of a condenser. It is at once evident that the above speculation confuses capacity with conductance—two definitely independent variables. We ask the authors to state in what group or subdivision their garnet belongs; what method and units were used in calculating the data of col. A, table I and their meaning; what was the temperature of the carrier roll and, finally, has any effort been made to investigate the effects of particle shape on distribution in the electrostatic field? S. B. Hudson—I have read this article with great interest. We have been engaged in research work on the principles of electrostatic separation in this laboratory for some time now, and our findings agree with those of the authors in many respects. The article shows evidence of careful and valuable research in the field of electrostatic separation. A "distribution analyser," very similar to that described in an earlier article by one of the authors," was incorporated in an inclined plate-type electrostatic separator designed and built in the Melbourne University laboratory in 1948 for investigation purposes.22 The actual splitting edges were machined from y! in. per-spex, and the paper hoppers were supported on linen thread immediately below the perspex dividers. These dividers fitted into machined slots in a framework to give accurate Yz-in. spacings. The hoppers (staggered) fed directly into a rack of test tubes, which is supported on a vertical pantograph arrangement. The rack was positioned with guides on the horizontal pantograph stand, and this ensured positive alignment when replacing the rack after making weighings. In later work, when much heavier feed rates were used, of the order of 30 to 40 Ib per in. per hr a rack fitted with rectangular metal containers and similarly aligned was used. Some work was done here on comparing the distributions of minerals when passed separately and when passed as a mixture, and it was found that there was quite an appreciable difference in the two results.= However, in our separator the particles do not pass down the plate in a single layer, and this difference is probably caused by collisions of one mineral particles with the other mineral particles. In most of the investigational work here, the change of the center point of the distribution is measured to establish the effect of a variable, such as voltage. Two minerals (zircon and rutile) have been studied rather exhaustively, and it was found that their distributions are very nearly normal. Owing to the sharpness of the distribution curves, the usual method of obtaining the mean or median was inaccurate, and was not used; instead the mean (also the median), calculated on the assumption of a normal distribution, was used to locate the center point of each distribution and proved satisfactory. The effect of polarity becomes very apparent in the plate-type separator where frictional charges play a very important part when using highly resistive minerals such as zircon. With rutile, a comparatively conductive mineral, polarity of the electrode has little effect. On the other hand, the magnitude of the voltage has a far greater effect on conductive than on resistive minerals. Shiou-Chuan Sun (authors' reply)—Thanks are extended to Drs. Ralston and Fraas for their keen interest in this paper. Their questions concerning coal dust,

Jan 1, 1951

By A. W. Asman

The problem of balanced design in a belt conveyor has been under constant study. Increased use of this method of transport in recent years has stimulated investigations. Conveyor belt, power transmission and other machinery should be adequate to perform the required service. All components of the belt conveyor should work as an efficient unit moving its load continuously between terminals. As part of a long term research project on belt conveyors, Hewitt-Robins established in 1954 a 2-year grant at Pennsylvania State University. Under terms of the grant, certain fundamental phenomena were to be studied. This work evolved a new concept in belt conveyor power determination based on the principle of belt tension. The Penn State contract furnished the foundation for later more detailed studies. From these most recent investigations, more specific factors were defined. In the final analysis, all of the data collected has been worthwhile in belt conveyor design. Units now operate over long distances at speeds of over 800 fpm. Long high lifts are commonplace. Where vertical or inclined shafts were the rule, belt conveyors receive first consideration. Although belt conveyors have moved forward rapidly, the future will see a tremendous gain in their use. The ideal test belt conveyor would be several thousand feet in length and have a maximum tension of at least 70,000 lb., Also, this conveyor should be horizontai and have a variable feed rate of more than one material. Data from such a field belt conveyor combined with laboratory information would permit accurate analysis. Since this type of conveyor was not available, many field locations were visited: An alternate would have been to erect a closed circuit system in the laboratory. Either method had limitations and the first was selected as having the best possibilities. Manufacturers of conveyor belting and machinery basically have been using the same horsepower formula. As more information became available from field installations, factors in the basic formula were modified. Practically all conveyors are designed today by adding together the horsepowers required to move the empty belt, move the load horizontally, and elevate the load. In selecting size of drive motor, the first two elements are relatively unimportant if the last is large. However, it is the combination of

Jan 1, 1961

By R. C. Weed

LEACHING in place at Cananea began in the 1920's on a limited scale. The first plants were small wooden boxes located underground in the Capote and Oversight mines, and output was low. Scrap iron was used as precipitant. Later several other small plants were started up. Leaching was limited entirely to underground areas, and production remained low. The Veta precipitating plant, the first on a larger scale, was put into operation in the late 1920's. After a period of intermittent service it was rebuilt in 1946 and has since been running continuously. In 1943 work on the Colorada open pit was started. This produced 40 million tons of waste, which averaged 0.20 pct capper. When these dumps had aged a few years it was decided to leach them on a large scale, and after considerable research and investigation a leaching system of pumps and pipelines was laid out and installed. At the same time the Ron-quillo precipitation plant was built to treat the copper solutions from the pit dumps. This plant started operation in March 1953. Leaching areas consist of the Colorada pit dumps and various mined out and caved underground stopes. The principal mineral, chalcocite, leaches readily in a weak sulfuric acid solution containing ferric iron. Enough pyrite is contained in the areas to produce an acid-ferric iron solution strong enough to dissolve the copper. Each area is studied to determine the best method of distributing the water over it and the best method of collecting the copper-bearing solutions after leaching has taken place. Water is generally distributed by means of ponds, ditches, and sprays. Spraying has proved the best method of spreading the water evenly; however, in a dry climate some water is lost through evaporation. After passing through the leachable material the solution is collected underground by dams and bulkheads and on surface by dams. An effort is made to give the maximum storage capacity practicable.

Jan 1, 1957

By D. E. Light, J. Y. Somnay

The use of alkyl acid phosphates and their alkali salts as collectors for the uranium minerals brannerite and uranothorite was investigated. In particular a detailed flotation study was carried out using isooctyl acid phosphate as a promoter. Tests were conducted on an Elliot Lake ore analyzing 0.11% U 3 O 8, chiefly as brannerite, and approximately 9% pyrite. Pyrite was floated with conventional reagents; the flotation of brannerite from the sulfide tails was studied to determine the optimum pH and pulp density to employ with isooctyl phosphate. 4 uranium recovery of over 90% was possible with a selectivity index of 8.95 when these tails were floated at 17% solids and a pH of 1.7. The effect of various modifying agents was also investigated. Aluminum sulfate, lactic acid and sodium silicate improved the selectivity of collection yielding selectivity indices of 9.5 to 11.5. Ferric chloride was found to be a depressant for brannerite. A Bancroft ore which analyzed 0.13% U308, and contained uranothorite was also amenable to flotation using isooctyl acid phosphate. The composite uranium concentrate obtained assayed 0.41% U308 at a recovery of 95.3% and a ratio of concentration of 3.24. Previous investigations1,2 on the flotation of brannerite, pitchblende, uraninite and/or uranothorite from various Canadian ores showed that fatty acids, including oleic acid and tall oils, petroleum sulfo-nates and alkyl acid phosphates, were suitable collectors for these uranium minerals. Additional work carried out by the Mines Branch in Ottawa on brannerite ores from the Elliot Lake area3 used the tall oil Acintol FA-1, or FA-2, as a collector. A recovery of 92% of the uranium in a concentrate which contained about 55% of the weight and assayed 0.2% U3O8 was obtained from a feed of 0.1% U308 in this latter instance. Eigeles et a1.4 carried out flotation on pitchblende ores and found that the most promising results were obtained with alkyl acid phosphates and their alkali salts as collectors. Within the homologous series of alkyl phosphates, the best flotation characteristics were displayed by isooctyl phosphates. Further, isooctyl phosphate derivatives were much less sensitive to hardness forming salts than oleic acid, enabling all flotation tests to be made with tap water. This paper describes tests employing alkyl acid phosphates for the flotation of ores containing brannerite and uranothorite. The flotation work described herein was conducted in conjunction with the development of a leaching process for Elliot Lake, Ontario, brannerite ores. In this process pyrite in the ore was concentrated by flotation and roasted, providing sulfur dioxide which was used to produce sulfuric acid in situ in a liquid-solid extraction process. A 40% conversion to sulfuric acid was indicated at a leaching temperature of 80°C; this resulted in the extraction of 93% of the total uranium after a contact period of 6 hr. A rotary kiln was used as a leaching vessel. Preconcentration by flotation was studied in conjunction with the above process to reduce the bulk of solids for leaching and, consequently, capital expenditure. The materials and methods used in this study will be described initially. Then the results will be presented in two sections, the first dealing with the brannerite ore, and the second with the uranothorite. Finally an economic analysis of the results on the brannerite material will be made. MATERIALS AND METHODS The two ore samples examined were from uranium-producing mines. One was a brannerite ore, a uraniferous quartz pebble conglomerate from the Elliot Lake area, which contained 0.11% U3O8, 4.3% S and 4.5% total Fe. Mineralogical studies of similar samples indicated the presence of quartz, microcline, sericite, pyrite and minor quantities of rutile, pyrrho-tite and zircon. The other ore was from a biotite-rich pegmatite in the Bancroft area; feldspar, quartz and biotite formed the major gangue components. Uranothorite was the principal uranium mineral; the head assay was 0.13% U 3 O 8. Preparation of Feed: The ores were crushed to —10

Jan 1, 1963

By J. H. Polhems, R. B. Brackin, D. B. Grove

THE heavy media separation process plays an outstanding role in the concentration of 4000 tons of zinc ore per day at the Mascot mill of the American Zinc Co. of Tennessee. Of the total tonnage, 72 pct is treated in the heavy media separation plant to reject 56 pct of the ore as a coarse tailing, which has a ready market. Concentrates from this separation are beneficiated further by jigging and flotation. Approximately 25 pct of the total zinc concentrate production is made in the jig mill. Jig tailings are ground and pumped to the flotation circuit where the balance of the production is made. Fig. 1 shows a generalized flowsheet of the mill. The Mascot ore is a lead-free, honey-colored sphalerite in dolomitic limestone, with lesser amounts of chert and some pyrite. A mineralogical analysis is given in Table I. After 10 years of successful operation with galena medium and treatment of nearly 10,000,000 tons of ore, a decision to convert to ferrosilicon was made early in 1948 because of the increasing price of galena and consequent high operating costs. The conversion was made on Nov. 6, 1948, and the results obtained since that time have shown remarkable improvement over those made with galena. The Table I. Mineralogical Analysis of Mill Feed, Pct Calcium carbonate 49.5 Magnesium carbonate 35.2 Iron oxide and aluminum oxide 1.5 Zinc sulphide 4.5 Insoluble 9.3 100.0 Table II. Comparative Data, Galena and Ferrosilicon Ferro- Diner-Gelenaa siliconb ence Operating costs per ton milled, ct. 21.21 9.12 12.09 Medium consumption per ton milled, lb 0.80c 0.15 0.65 Reagent consumption per ton milled, lb 0.45 0.02 0.43 Tailing assay, pct Zn 0.310 0.297 0.013 Concentrate. oct Zn 12.08 10.33 1.75 Heavy medla ieparatlon recovery. pct 89.38 90.22 0.84 Mill feed rate, tons per hr 153 166 13 Heavy mesa separation feed rate. tons per hr 100 10 0 Tons milled per heavy media separation man shift 350 620 270 Mill feed to coarse tailings, pct 51.0 56.7 5.7 Lost mill time, pct 5.6 5.0 0.6 Power consumption, kw-hr per ton 2.06 1.92 0.14 a 1947. " First 6 months of 1950. c Net consumption after deducting credit for reclaimed waste galena. Consumption of new galena was 1.320 lb per ton milled. For entire life of galena operation, a credit of 40 pct of the value of the new galena added was realized from the sale of waste galena. comparisons given in this report cover the first 6 months of 1950 as representing the ferrosilicon operation, and the year 1947 as representing the galena operation. This was the last full year in which galena was used exclusively and is representative of the best work done during the 10 years of operation with this medium. After only 2 years' operating experience, with ferrosilicon and treatment of 1,807,585 tons many advantages have been revealed and are summarized in Table 11. Development Prior to the introduction of the heavy media process, all the mill feed was crushed through 5/8 in. and treated by jigging. A finished tailing assaying 0.66 pct Zn was made on rougher bull jigs, and cleaner jig tailings were ground for treatment by flotation. The first test work on the sink-and-float method of mineral beneficiation was carried out at Mascot in 1935, using a 3-ft cone and galena medium for batch tests. The following year a 6-ft cone was installed for pilot-plant work. This unit became a part of the mill circuit on March 1, 1936, and handled a gradually increasing tonnage in the next 2 years as the process developed to the point where it could treat all the + 3/8-in. material in the mill feed. Coarse jigging was then discontinued on March 1, 1939, and all coarse tailings have been made by the heavy media separation plant since that time. Feed Preparation: The original feed preparation plant consisted of a drag washer followed by two 4x10-ft Allis-Chalmers washing screens. A surge bin and two additional 5x12-ft AC washing screens were added in 1943. Use of primary and secondary washing screens was found essential to provide the cleanest possible feed for the cone and thereby avoid excessive contamination of the galena medium. Improved washing was obtained by replacing the drag washer with a 7x20-ft Allis-Chalmers scrubber, shown in Fig. 2, which has been in service since May 1944. Throughout the life of the galena operation, delivery of extremely muddy ore to the mill overloaded the medium cleaning system, and it frequently was necessary to cut off the feed and clean the medium for several hours until its normal viscosity had been re-established. The cleaning circuit

Jan 1, 1952

By R. G. Knickerbocker, W. E. Brown, G. J. Vahrenkamp, M. M. Fine

THE results of a research and development laboratory and pilot-plant mineral-dressing investigation by the Bureau of Mines of the complex copper-lead-cobalt-nickel ores of southeast Missouri are reported in this paper. This investigation, in cooperation with the National Lead Co., showed that separate concentrates of lead, copper, and cobalt-nickel could be made by fine grinding and flotation. An industrial plant built subsequently by the St. Louis Smelting and Refining Div. near Fredericktown, Madison County, Mo., produced such concentrates during the war years. Present commercial production at the mill is limited to lead and copper concentrates, but research to permit economic utilization of the cobalt and nickel values is continuing. The complex sulphide deposits near Frederick-town in Madison County, Mo., have an operating history that dates back to the years preceding the Civil War. Perhaps these operations can best be summarized in terms of output which are given by Tyler: "Previous to 1900, the Mine la Motte Co. shipped cobalt ore to Swansea and in 1903 produced 120,000 pounds of oxide in a Missouri plant. Another cobalt refinery erected in 1906 by the North America Lead Co. at Fredericktown, Mo., operated intermittently until 1910. In 1909 it produced 83,384 pounds of cobalt oxide, 328,403 pounds of nickel, 8,314 tons of nickel-cobalt concentrates, 600 tons of copper concentrates, and 1,353 tons of lead concentrates. This plant was purchased at a forced sale in 1915 by the Missouri Cobalt Co., which produced a certain amount of oxide in 1918 and in 1920 recovered 102,410 pounds of cobalt oxide and cobalt hydrate, valued at $262,810." According to Buehler,2 "The ore, which consists of a complex mixture of the sulphides of cobalt, nickel, copper, lead, and iron, occurs at the contact of the pre-Cambrian porphyry and the overlying Lamotte sandstone and Bonneterre dolomite." The source of most of the cobalt and nickel is siegenite, a member of the linnaetite series of minerals. Its composition is variable, and although the chemical formula may be given as (CoNi),S,, it usually contains both iron and copper. The siegenite is characterized in this instance by an intimate association with chalcopyrite, and to a lesser extent with iron sulphides. Galena, on the other hand, while dis- seminated throughout the ore, is not closely bound to other constituents. Gangue is composed essentially of dolomite and quartz. Petrographic examination showed that fair liberation of sulphides and gangue could be attained at — 100 mesh. Complete liberation of micronsize inclusions, particularly of sulphides in each other would not have been possible at —400 mesh. Beneficiation Studies Many samples of the southeast Missouri ore were tested, but most of the work was confined to five samples of churn drill cuttings weighing 15 tons. The samples were reportedly representative of five sections of the deposit and were tested individually and as a composite. Only the results of the tests on the composite, which contained 1.28 pct lead, 2.03 pct copper, 1.06 pct nickel, 0.87 pct cobalt, 3.57 pct iron, 15.4 pct lime, and 23.9 pct silica, are presented in this report. In view of the rather complex and intimate association in the mineralogical composition of the ore, flotation was the logical choice in the selection of a beneficiation process. Other mineral dressing unit operations were considered as an adjunct to flotation, but none were beneficial. Many tests were made on these samples to determine the flotation characteristics of the mineral siegenite. As a result of this work, it was noted the siegenite was intermediate in floatability between chalcopyrite and galena on one extreme and iron sulphides on the other. The difference in floatability between siegenite and the iron sulphides was not very great and, in general, the same reagents that depressed pyrite and marcasite would also, in slightly larger quantities, depress siegenite. This tendency for siegenite

Jan 1, 1952

By T. B. Counselman

"C" VERYONE who has to calculate cyanide circuits, -¦-' using either thickeners, filters or both, realizes the headaches involved in solving a set of simultaneous equations. When you calculate a large number of these, based on experimental data, trying to arrive at the best flowsheet, you can spend days at the task. T. B. COUNSELMAN, Member AIME, is Manager, FluoSolids Division, The Dorr CO., Stamford, Conn. AIME New York Meeting, Feb. 1950. TP 2783 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Aug. 23, 1949. Herewith is presented a short method, equally accurate with the standard simultaneous equation method, and much quicker. Here for instance is the standard simultaneous equation method, for 100 tons per day of $10 ore, with the conditions assumed as below (fig. 1). Conditions Assumed a. 100 tons of ore per day crushed in cyanide solution. b. Discharge from all thickeners with 50 pct moisture. c. $10 value dissolved per ton of ore. d. 50 pct in mill and 50 pct in agitators. e. 400 tons of solution from thickener V precipitated to $0.02. f. Agitation with a dilution of 2 of solution to 1 of solids. g. Let V, W, X, Y and Z represent value in dollars per ton of solution discharged from the resvective thickeners. Equating Out of and Into Each Thickener 1. lOOVplus 400V = 500W plus (0.50 X $10 X 100) 2. 1OOWplus 600W= 500X plus lOOW plus (0.50 X $10 X 100) plus lOOV 3. lOOXplus 500X = 1OOWplus 500Y 4. 1OOY plus 500Y = 1002 plus 1OOX plus (400 X 0.02) 5. 100 Zplus l00Z = l00Y plus 100 tons of water (value $0.00) Simplifying: 1. v = W plus 1.00 2. W = X plus 1.20 3. X = Y plus 0.24 4. Y = 0.2Zvlus 0.064 5. 2Z= Y

Jan 1, 1951

By T. B. Counselman

"C" VERYONE who has to calculate cyanide circuits, -¦-' using either thickeners, filters or both, realizes the headaches involved in solving a set of simultaneous equations. When you calculate a large number of these, based on experimental data, trying to arrive at the best flowsheet, you can spend days at the task. T. B. COUNSELMAN, Member AIME, is Manager, FluoSolids Division, The Dorr CO., Stamford, Conn. AIME New York Meeting, Feb. 1950. TP 2783 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Aug. 23, 1949. Herewith is presented a short method, equally accurate with the standard simultaneous equation method, and much quicker. Here for instance is the standard simultaneous equation method, for 100 tons per day of $10 ore, with the conditions assumed as below (fig. 1). Conditions Assumed a. 100 tons of ore per day crushed in cyanide solution. b. Discharge from all thickeners with 50 pct moisture. c. $10 value dissolved per ton of ore. d. 50 pct in mill and 50 pct in agitators. e. 400 tons of solution from thickener V precipitated to $0.02. f. Agitation with a dilution of 2 of solution to 1 of solids. g. Let V, W, X, Y and Z represent value in dollars per ton of solution discharged from the resvective thickeners. Equating Out of and Into Each Thickener 1. lOOVplus 400V = 500W plus (0.50 X $10 X 100) 2. 1OOWplus 600W= 500X plus lOOW plus (0.50 X $10 X 100) plus lOOV 3. lOOXplus 500X = 1OOWplus 500Y 4. 1OOY plus 500Y = 1002 plus 1OOX plus (400 X 0.02) 5. 100 Zplus l00Z = l00Y plus 100 tons of water (value $0.00) Simplifying: 1. v = W plus 1.00 2. W = X plus 1.20 3. X = Y plus 0.24 4. Y = 0.2Zvlus 0.064 5. 2Z= Y

Jan 1, 1951

By T. B. Counselman

C. G. McLachlan—In the foregoing paper the author has presented a very neat method for calculating the solution recovery for a countercurrent flowsheet. He has, however, based his calculations, as he states, on the assumption that "the concentration of dissolved value must be exactly the same in the overflow and underflow of any thickener." This assumption, which is usually made regardless of the type of calculation used, is "not strictly correct because diffusion of the dissolved gold in the solution associated with the solids as it progresses through the countercurrent system is far from being instantaneous, with the result that on a ton for ton basis, the assay of the solution in the thickener underflow is higher than in the corresponding thickener overflow. An agitator introduced into the flowsheet between thickeners will reduce, but not entirely eliminate, this difference. In practice we have found that the best way to take care of this condition— in addition to the use of one or more intermediate agitators—is to circulate 20 pct or 25 pct more solution than called for by the cheoretical calculation. This presents no difficulty provided that adequate thickener capacity is provided in the original design of the countercurrent circuit.

Jan 1, 1951

By T. B. Counselman

C. G. McLachlan—In the foregoing paper the author has presented a very neat method for calculating the solution recovery for a countercurrent flowsheet. He has, however, based his calculations, as he states, on the assumption that "the concentration of dissolved value must be exactly the same in the overflow and underflow of any thickener." This assumption, which is usually made regardless of the type of calculation used, is "not strictly correct because diffusion of the dissolved gold in the solution associated with the solids as it progresses through the countercurrent system is far from being instantaneous, with the result that on a ton for ton basis, the assay of the solution in the thickener underflow is higher than in the corresponding thickener overflow. An agitator introduced into the flowsheet between thickeners will reduce, but not entirely eliminate, this difference. In practice we have found that the best way to take care of this condition— in addition to the use of one or more intermediate agitators—is to circulate 20 pct or 25 pct more solution than called for by the cheoretical calculation. This presents no difficulty provided that adequate thickener capacity is provided in the original design of the countercurrent circuit.

Jan 1, 1951

By H. A. Hoffman

Competition from an all-flotation plant, with demonstrated economies and efficiencies, plus a change in smelting contract and introduction of improved cyclones lead to conversion from gravity-flotation. Detailed descriptions are given of equipment installed and procedures used at St. Joseph Lead's Federal mill. Future plans include further quality control by instrumentation in all aspects of the mill operation. Also planned are improvements in materials handling systems. Competitive conditions will continue to dictate improvements and changes in the mill flowsheet. The advent of all-flotation mills in the Lead Belt was introduced by the Indian Creek mill in late 1953. This modern and efficient plant then became a pattern for the other mills in this area to re-evaluate their circuits in an effort to develop flowsheets that would improve operating conditions and metallurgy. The Indian Creek mill demonstrated that all-flotation would require considerably less operating and maintenance labor than the combination of tabling and flotation that was common in the Lead Belt at that time. Two significant changes occurred in recent years to allow all-flotation to be seriously considered in other Lead Belt mills. One was a change in the smelting contract which did not require gravity concentrate. Another was the development of cyclones which provided classification of the flotation feed in a very small space. Since the Federal mill was the largest concentrating plant in the Lead Belt it was felt that the greatest savings could be made by investigating the all-flotation possibilities at this mill. Interest was stimulated by the milling and ore dressing Depts. to determine if this mill would be converted without undue cost, provided the metallurgy could be improved. Accordingly, laboratory tests were initiated to learn what metallurgical benefit could be derived. Numerous tests were run which indicated that grinding to all-flotation would improve the tailing by as much as 0.02 pct Pb. This was quite significant when multiplied by the tonnage of ore treated. Projecting the added cost of power for extra grinding and flotation, and the additional flotation reagents required, plus additional new equipment that would have to be purchased, it would still add up to a considerable saving. The all-flotation mill would reduce man power by some 30 pct, and make a metallurgical improvement. Operating costs could be reduced bv 2$ per ton. The metallurgical improvement would amount to 4$ per ton, for a total of 6$ per ton of ore milled. An estimate of the cost of the conversion was set at $250,000. With the savings as estimated this could be paid off in about two years. On paper it therefore seemed attractive enough to justify a mill test. MILL TESTING One section of the Federal mill was made available for use as a separate test circuit. Cyclones and density controllers were borrowed from the Viburnum and Indian Creek mills. Denver flotation cells were obtained from the Desloge mill that was then shut down. Two 10-cell groups of Fagergren flotation machines, consisting of eight roughers and two cleaners in each bank, were segregated from the remainder of the plant. The 9x12-ft rod mill was continued as a primary grinding mill. The two 6%x12-ft mills on this section were converted from rods to balls, and the speed increased to 22.0 rpm. Cyclones were installed to close the circuit of these ball mills. The ground pulp from each of the ball mills was fed separately to its own bank of flotation cells so that two completely separated test circuits could be run in parallel. This test circuit operated on a three-shift basis from Dec. 22, 1959 to Mar. 21, 1960. The first few weeks were occupied in developing conditions for proper metallurgy and trouble-free operations. Results were rather erratic but as the ore dressing laboratory and the mill operators became more familiar with conditions they were able to obtain expected results, which could be duplicated day after day. The last five weeks of testing indicated the best results and are tabulated in Table I. During the first two weeks a Denver unit cell was used on No. 1 flotation circuit. Even though it recovered almost one half of

Jan 1, 1962

By R. J. Brison, O. F. Tangel

The development of a new method of mineral separation was sponsored by the International Salt Company, which requested Battelle Institute to investigate means for improving the quality and appearance of rock salt from the Company's Detroit mine. Although developed specifically for removing impurities from rock salt, the general method may be applicable to other separation problems. The principal impurities in rock salt from the Detroit mine are dolomite and anhydrite which represent 2 to 5 pct of the weight of the mined salt. In the size range from 1/4 to M in. (the range of primary interest in this project) the impurities are only partially liberated from the halite in normal production. Further size reduction to improve the liberation of impurities is not practicable in view of the market requirements for the coarse grades of rock salt. Laboratory separations in heavy liquids showed that, to improve the quality and appearance of the rock salt substantially, it would be necessary to remove not only free gangue particles but also a large proportion of the locked-in particles. Because rock salt is an inexpensive commodity, a low-cost process was required. Gravity methods were, of course, considered. The heavy-liquid separations indicated that a split at an effective specific gravity of 2.2 to 2.3 would be required. (The specific gravity of pure halite is 2.16.) Heavy-media separation was investigated but had the disadvantages that it was necessary both to operate with saturated brine and to dry the cleaned salt, and that the cleaned salt was darkened by the magnetite medium. Air tabling was tried but did not give the desired separation. It soon became apparent that established methods would not provide a satisfactory solution and work was undertaken on the development of a new process to solve the problem. PROCESS DEVELOPMENT Preliminary Experiments: At the start of the investigation, an analysis of the problem indicated that the diathermacy of rock salt—that is, its ability to transmit radiant heat—might form the basis for an efficient separation process. Under this theory, the impurities might be selectively heated by radiant heat. The particles could then be fed over a belt coated with a heat-sensitive substance so that the warm impure particles would adhere preferentially to the coating. After the initial experiments, made by heating the rock salt with an infrared lamp and separating the product on small sheets of resin-coated rubber, proved encouraging, a small continuous separation unit was set up. This comprised 1) a simple heating unit consisting of a vibrating feeder covered with aluminum foil and an infrared lamp mounted above the feeder and 2) a separation belt 6 in. wide and 36 in. long. A sketch of the device is shown in Fig. 1. Results with this apparatus confirmed the fact that a good separation was possible. It was apparent, however, that a considerable amount of experimental work would be needed to develop the scheme to a practical and economical process. The Process: Basically, the process consists of two main steps: 1) selective heating by radiation and 2) separation of the heated particles on a heat-sensitive surface. Because neither of these steps had previously been utilized commercially in mineral processing, it was necessary to do basic research on both aspects. Factors studied in the investigation included type of heat source, design of heating unit, design of separation belt, selection of heat-sensitive coating, removal of heated particles from the belt, contact between particles and coating, and maintenance of the heat-sensitive surface. Part of the experimental work was carried out on a small-scale unit consisting of the 36x6 in. belt and auxiliary apparatus, and part on a larger unit. For simplicity, discussion of work on both of these units is grouped together. SELECTIVE HEATING Radiant-Heat Source: The essential requirements for a radiant-heat source were 1) that the radiant heat be in a wave length range which is effectively absorbed by the impurities but not absorbed appreciably by the rock salt and 2) that it be dependable, practical, and economical. Selection of a heat source of suitable wave length range was one of the first considerations. It is well known that pure halite is highly transparent to radiant energy in wave lengths from 0.3 to 13 microns. However, the available data on infrared transmission by dolomite and anhydrite, particularly in the range below two microns, were not complete enough to serve as a reliable basis for selection of a heat source. Although it may have been possible to obtain sufficient data on infrared transmission and absorption to enable one to select the best heat source, a more direct procedure was used. This consisted simply of exposing the crude rock salt to each of several types of radiant-heat source on the small continuous separation device. The heat sources investigated, approximate source temperature used, and calculated wave length of maximum radiation are tabulated in Table I. Of the two types of tungsten-filament lamps investigated, both the short wave length photoflood lamps and the longer wave length infrared lamps were satisfactory from the standpoint of selectivity

Jan 1, 1961

By O. E. Palasvirta

Several drum-type, low-intensity dry magnetic separators have recently been described 1,2,3,4 all of which share at least two generic similarities. The magnet consists of a multiplicity of poles of alternating polarity and, the speed of the drum shell, relatiw to the magnet, is such that the magnetics are agitated through rotation of the magnetic particles around their individual axes. The speed of the drum relative to the magnet (differential speed) determines the frequency of the field reversals. For effective agitation this frequency generally has to be above a specific value, which varies according to the mean particle size of the magnetics. With separators of small diameter, say less than 11/2 ft., a sufficient differential speed for treatment of fine magnetite can be attained only by rotating both the magnet and the drum separately; separators of large diameter may attain sufficiently high differential speeds with stationary magnets. Separators of large diameter thus have the advantage of simplicity of design. The capacity of a separator of this type is directly proportional to the speed of the shell. As the centrifugal force opposing the magnetic force increases at a slower rate than the peripheral speed of the shell when drum diameter is increased, large drums can be operated with higher peripheral speeds. Thus they have a higher capacity than smaller drums. Separation is also a statistical phenomenon, i.e. the grade of concentrate improves as the number of agitations or rotations of the magnetic particles per separation increases. As separators of large diameter have space for a large number of magnets they are best suited for production of optimum grade concentrate. The statements above are general clues For separator design. In the following an expression IS derived showing a more exact relationship between the pertinent separator variables. As both magnetic attraction and centrifugal forces are directly proportional to the volume of the affected bodies, complications in mathematical analysis, arising from variable particle size, are avoided by considering the force per unit volume of affected material. For the magnetic force per unit volume Runolinna (3) gives the following equation: V " 1 + N(ur, - 1) uo dr (1) where ur and uo stand for the relative permeability and the permeability of vacuum, respectively, N is the demagnetization factor, B the intensity of the magnetic field, and r the distance from the face of the magnetic pole. The permeability of a specific magnetite ore can generally be considered as a constant. For ground, liberated material N can also be treated as a constant, and the first simplification can be made by writing ur-1 1=k1. 1 + N(ur - 1) uD k1. Fig. 1 shows the field intensity curves for two U-magnets of different size, but of roughly similar geometry of cross-section as plotted against the distance from the plane of the pole faces. The curves

Jan 1, 1961

By Pablo Hadzeriga

Great Salt Lake brine was subjected, in laboratory scale, to conditions simulating solar evaporation. Solid phases and the variation in composition of the liquid phase throughout the potassium salts crystallization zone are presented. These experimental results are discussed and compare with stable equilibrium data of the quinary system Cl-, SO4, Na+, K+, Mg + + and H2O. The Great Salt Lake brine in Utah is a vast reserve of several potential chemicals which has attracted much attention in the recent past. From the standpoint of an ore reserve, it may be said that the lake contains a low grade ore which has to be concentrated by evaporation of water. Because of the fairly favorable weather conditions in the Great Salt Lake area, it seems reasonable to consider solar evaporation as an economical means of concentration. From the view of phase chemistry, the main constituents of Great Salt Lake brine make it belong to the five component system: C1-, SO4=, Na+, K+, Mg++ and H2O. Because of their low concentration, the presence of other elements may be regarded as having little effect on the equilibrium of the five main components. Equilibrium data for the quinary system of our interest has been available for many years. Using these data, it is possible to calculate how the brine of the lake will concentrate when subjected to solar evaporation. In practicing solar evaporation, the weather factor and supersaturation conditions cannot be ignored, but their introduction into calculations using equilibrium diagrams is almost an impossible task. With the experimental work presented in this paper, an attempt is made to compare the stable equilibrium data of the quinary system as applied to the case of the Great Salt Lake brine and the dynamic equilibrium when that brine is subjected to continuous solar evaporation. Sea water brine belongs to this quinary system when only the major constituents are considered. The study of this system has been carried on since the last decades of the nineteenth century. The reason for these intensive studies was to give a geological explanation to the underground potash deposits. AS a result, many theories were presented trying to relate the equilibrium data with the geological formation, paragenesis and metamorphism of these deposits. Two main approaches were followed for the equilibrium studies of this quinary system. One of them, and the best known to us, was the school started in Germany in the last century by J. H. Van't Hoff. The other one, which started about the same time, was the school of N. S. Kurnakov in Russia. Both approaches were different and as a consequence generated opposite theories about the geological formation of potash deposits. The Van't Hoff school studied the quinary system of sea water under careful laboratory control, resulting in stable phase diagrams. During these studies, Van't Hoff recognized the difficulties in crystallizing some double salts.' ". . . . we unexpectedly found that some compounds whose formation in the solutions under investigation was possible at 25°C, nevertheless totally failed to put in an appearance. These, not to mention the calcium salts, were leonite, MgSO4 . K2SO4 4 H2O; kainite, KC1-MgSO4 . 3 H2O; and kieserite, MgSO4 .H2O. Even exceedingly slow crystallization with the addition of the compounds themselves did not remove the condition of supersaturation in the case of these bodies." However, those salts (leonite, kainite, and kieserite) were included in the stable phase diagram at 25°C as shown in Fig. 1 using D'Ans data.3 The figure also shows the sequence of crystallization of the Great Salt Lake brine according to the stable diagram. In later studies by the Van't Hoff school, especially by D'Ans,3 Karsten,4 Autenrieth,5,6 attention

Jan 1, 1968

By R. J. Linney

Reserve Mining Co. produces 5 million tons of iron ore pellets per year. The finished product runs about 65.50 pct Fe, with 7.75 to 8.00 pct SiO2. Less than 12 pct arrives at the blast furnaces smaller than % in. This plant was originally designed to produce 3.75 tons annually—not 5 million. To handle the increased tonnage a fifth fine crushing line was added, and some very minor changes were made. When production can be stepped up 33 1/3 pct without extensive additions or changes—within the short period of a year—basic engineering has been sound. Operators of the plant consider it the best design job they have seen or heard of, and competent authorities have said that start-up was quicker and easier than for any previous large-scale minerals beneficiation plant. One big advantage is the elbow room allowed. Generally speaking, there is an eternal conflict between operator and designer over space. The designer has his costs per cubic foot, and he is keenly aware that if he permits the number of cubic feet to rise unnecessarily, total costs will be too high. But this was a new operation, and on a large scale. There were no proven standards to go by. Realizing this, Reserve's engineers provided ample and efficient access to machinery, allowing extra space for unanticipated situations. In three years of production this planning has paid off handsomely. Gravity Flow Saves Time and Money: This plant is on a rocky hillside, the car dumper about 270 ft above Lake Superior. In locating buildings and placing machinery the designers mad= good use of gravity flow, at considerable financial savings. In the concentrator building from rod mills to filters, gravity carries the material between all stages of the operation except one—the closed circuit pumping of concentrate between cone classifiers and ball mills. Since no machine is located directly above any other, an observer can stand on the truck ramp adjacent to the offices and laboratory and see practically every phase of activity in the huge building. The first mill section started in August 1955. After a day's run several modifications were made on auxiliary chutes and launders where spillage had occurred. This procedure was followed for the first

Jan 1, 1960