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The Cadia mine is shifting from the open pit operation to the Cadia East underground block cave. The resulting increase in milled tonnes and the decrease in fines production through shifting to block caving require an increase of over 25% in installed power. This expansion is being accomplished via the installation of up-front crushing and HPGR milling of the ore, with the SAG mill receiving a secondary crushed feed. This design, criteria and rationale for selection is presented. Over the coming years, the mine will further expand capacity. The target is to accomplish this with minimal addition of installed power or use of fresh water. This presents a challenge that epitomizes the future constraints on our resource dependent and intensive industry. The approach and proposed options to this longer-term expansion are also presented.

Jan 1, 2012

By Walter Estrada, Fernando Valencia, Jaime Luyo

La gasificación por plasma es un proceso térmico a alta temperatura en el que los desechos se descomponen en elementos fundamentales y compuestos simples en un entorno anaeróbico. El sistema presentado en este artículo actúa como un digestor al transformar las moléculas de carbono y otras sustancias en compuestos menos contaminantes. Se utilizó una mezcla de cianuro disuelto en agua como muestra de ensayo —procedente de aguas de una mina—, que se sometió al componente radiante de plasma, y se estudió la energía generada actuando sobre la purificación y transformación de las muestras.

Sep 30, 2019

By B. W. Hester

Report of discussion at January, 1971, general meeting (Chairman: M. J. Cahalan, President) of the paper published in Transactions/Section B (Applied Earth Science), vol. 79, May, 1970, pp. B60-B67, and contributed remarks. Comments were made by Taylor J.T.M., James C.H., Lethbridge R.F. and Heath K.C.G.

Dec 1, 1971

By Paul H. Johnson

Heap leaching is a simple means of recovering copper from oxide and oxide-sulfide ores. Although this method requires relatively low capital investment, it suffers from low copper recoveries and relatively high acid consumption. The objectives of this study were (1) to determine what factors influence these shortcomings, and (2) to establish a basis of correlation. between laboratory and field testing. The principal methods of leaching copper ores at the present time are heap, dump, in-place, and vat. A number of references are presented in the bibliography on the theoretical and practical aspects of these methods. In the heap-leaching method, crushed or uncrushed oxidized copper ore is placed on a prepared drainage pad. Acid leach solutions are dis¬tributed over the top of the heap, caused to percolate downward through the ore bed by gravity, and are drained from its base. The leach duration is generally a matter of a few months. Dump leaching is used to recover copper from very low-grade oxide-sulfide pit-waste materials. Here again, leach solutions are, percolated downward through a heap. However, in this method, little or no acid is added to the leach liquors, prepared drainage pads are not used and the leaching cycle is generally measured in terms of several years. Leaching agents are usually autogenously generated by the breakdown of sulfide minerals within the dump.

Jan 1, 1970

By Paul C. Merritt

Samuel Hopkins, an 18th century inventor from Philadelphia, has been little noted nor long remembered by History, but it was he who on July 31, 1790, obtained what no other man can ever achieve -the first patent issued by the Government of the United States. Signed by President George Washington, the patent guaranteed to Hopkins the protection of the American Government for his "improvement, not known or used before, . . . in the making of Potash . . . by a new Apparatus and Process." The patent itself is no longer important, but it serves as a milestone in the development of an industry that is now rapidly moving to fulfill its role as a mineral to feed the world. Potash ranks as one of the oldest products of Man, having been used as a condiment by the habitants of Gaul a thousand years before Athens achieved its Golden Age of culture. By Hopkins' time, the uses of potash had changed considerably, going primarily into glass-making, tanning, soap, and explosives, yet the same sources and techniques of manufacturing potash used during the earlier 2000 years-that is, leaching of plant ashes in iron pots with water, followed by water evaporation and crystallization of the potassium salts-were still being used.

Jan 10, 1966

By J. C. Kingston

Introduction The sliding-form method of erecting concrete structures was first devised and used in the early part of this century by the Sante Fe Railroad Company for the erection of a grain elevator at Santa Fe, New Mexico. A somewhat related principle has been employed for a number of years in shaft lining work and, to our knowledge, was perfected in Butte, Montana, but until the building of the concrete headframe at the Jeffrey mine of the Canadian Johns-Manville Company, Ltd., at Asbestos, Que., the method, as applied to surface structures, was quite foreign to the mining industry in Canada. The current shortage and cost of structural steel shapes has no doubt contributed more than any other single factor to the investigations that have led to the use of reinforced concrete as a headframe building material. Mining men in western Quebec and northern Ontario have shown a great deal of interest in the design and construction of the headframe at the East Sullivan mine in Bourlamaque township, Que., and we have had many queries regarding the methods used and, in particular, for an explanation of the sliding-form principle. This article is written in an attempt to set forth the conditions which were to be met; to explain the construction methods used, and to discuss the economy of concrete headframes as compared with the conventional steel structures with which the mining fraternity is familiar.

Jan 1, 1949

By P. M. Ambrose

This report describes an accident, which took place in the Chemical Corrosion Laboratory at the College Park, Md., Station of the Federal Bureau of Mines. The accident occurred while Walter L. Acherman, chemist, was performing a routine manipulation during a corrosion experiment involving commercially pure titanium and red-fuming nitric acid. There was a sudden and very violent reaction, and a large quantity of acid was thrown in the face and over the body of Acherman, causing serious injury. L. B. Golden, chief of the Chemical Corrosion Section, was exposed to nitric acid fumes during the ensuing rescue and first-aid operations and died on January 25, 1954. According to the medical report, the direct cause of Mr. Golden's death was pulmonary edema and confluent bronchopneumonia; the contributary cause, was inhalation of fumes from red fuming nitric acid. The following employees of the College Park Station contributed to this report by supplying information on the accident and assisted in the rescue operation: L. B. Golden, chief, Chemical Corrosion Section; T. E. Green, chemist; Robert Warf field, chemist; L. A. DeVivo, laboratory mechanic; D. W. Chalkley, laboratory mechanic; E. H. Neagle, laboratory welder; and R. L. Heath, laboratory general mechanic.

Jan 1, 1955

By Larry M. Southwick

Considerable process and equipment development and analysis were conducted during design of Freeport Sulfur's laterite nickel facilities in the late 1950's. These covered ore concentration at Moa Bay, Cuba, and concentrate refining at Port Nickel, Louisiana. Both facilities utilized acid pressure leaching and product refining steps that were similar to those later used elsewhere in the laterite processing industry. Unfortunately, developments in Cuba removed the Moa Bay facilities from Freeport control. This deprived the project of a start-up that would have allowed thorough resolution of difficulties and a quicker attainment of production capacity. This meant lost opportunity to positively prove and complete the technology when optimum resources and staff were available. This paper reviews the original rationale of how design was developed via laboratory, bench scale and pilot studies by Freeport Sulfur. Some later operating problems will also be discussed. This case is unusual, but the lessons learned may be instructive in the application of difficult and complicated technologies, as well as the need to understand and retain access to the original development and design activities Keywords: Moa Bay, Freeport Sulfur, Port Nickel, acid pressure leaching, pilot plant, pachuca, Chemical Construction, pressure letdown, cobalt, titanium

Jan 1, 2004

By Charles D. Litton

This U.S. Bureau of Mines report details the results of a series of large-scale experiments where small coal fires were used to ignite the conveyor belt at air velocities ranging from 0.76 to 6.1 m/s. In the tests, electrical strip heaters imbedded within a pile of coal were used to heat the coal to a point of flaming ignition. The flaming coal subsequently ignited the conveyor belt located approximately 5 to 10 cm above the coal pile. During the tests, temperature, CO, and smoke levels were continuously measured in order to determine both alarm time and level as the fire intensity progressed through the stages of smoldering coal, flaming coal, and flaming coal plus flaming belt. Analysis of the data leads to certain conditions of air velocity and sensor alarm levels that are required for early detection of conveyor belt entry fires. Two nomographs are presented, which define sensor alarm levels and sensor spacings as a function of belt entry cross-sectional area and belt entry air velocity.

Jan 1, 2010

By Brij M. Moudgil

Florida produced more than eighty percent of domestic phosphate during 1974. This amounted to more than thirty-five million short tons. Ninety percent of this production was used in the production of fertilizers. Other uses of phosphate include food preservatives, fabric dyes, optical glass, photographic films, soaps, plastics, pharmaceuticals and animal feed ingredients. Marketable phosphate rock projections for the nest decade forecast a very healthy growth for the Florida phosphate industry. Associated with the expansion, however, are the corresponding problems of water conservation, air and water pollution and land reclamation. Recognizing mining as an interim land use, the industry is presently engaged in a very active reclamation program. The proximity of the central Florida deposits to rapidly developing urban areas makes it important to undertake successful and well planned reclamation programs. Water use and conservation is of vitil importance to the industry. Water withdrawal from existing aquifers and discharges into drainage systems are closely regulated by state agencies. Water is stored and recycled to maximize usage. In some instances water is even returned to major aquifers through a series of recharge or connector wells. The phosphate industry has spent over 200 million dollars in the past decade for research, development and implementation of pollution abatement and water conservation systems.1 The following discussion outlines various phases of phosphate mining in central Florida today, with emphasis on research in waste disposal techniques and progressive land reclamation programs.

Jan 1, 1975

By Robert L. Clayton, Arthur Baker

Two massive sulfide zinc-copper ore bodies are in quartz-sericite schist (probably formed by regional metamorphism of sediments) and andesite of the Precambrian Yavapai Series, on opposite sides of a large rhyolite mass. The Copper Queen ore body is conformable with bedding of its schist host rock, but the Old Dick ore body probably crosscuts the strata. A lens of low-grade mineralization in chlorite schist, associated with the Old Dick ore body, crosscuts through 150 feet of strata. The ore bodies are almost entirely enclosed by premineral faults with less than 150 feet of displacement. The minerals of the ores are pyrite, sphalerite, and chalcopyrite, with minor arsenopyrite and galena and small amounts of intimately associated chlorite, sericite, and calcite. Old Dick sphalerite-chalcopyrite intergrowths have mutual boundaries, and sphalerite grains contain abundant exsolved chalcopyrite. Copper Queen sphalerite-chalcopyrite intergrowths have extremely ragged boundaries, the sphalerite contains no exsolved chalcopyrite, and the ore overall has a schistose texture. The pre-mineral fault on the hanging wall of the Copper Queen ore body has had 100 feet of post-mineral displacement. Wall rocks of the Old Dick ore body locally have been sericitized. Andesite and quartz-sericite schists for at least several hundred feet around both ore bodies have been chloritized erratically and contain traces of copper.

Jan 1, 1968

By Jorge Alberto Soares Tenório, Denise Crocce Romano Espinosa, Morgana Rosset, Bárbara da Rocha Pereira

High-purity alumina finds application in various electronic industry sectors, with approximately 100 million tons of this compound being produced worldwide. Alumina is acquired through the mining process of bauxite, a rock composed of a mixture of minerals such as gibbsite, kaolinite, hematite, and either anatase or rutile. In 2020, both bauxite and alumina were included in the list of critical raw materials. Due to the demand for high-purity alumina production, the development of a methodology to remove impurities present in the leach liquor from bauxite processing by solvent extraction (SX) followed by precipitation has been widely studied. This work aims to obtain high-purity alumina from bauxite by a solvent extraction process followed by a precipitation step from the sulfur leach liquor obtained by the thorough dissolution of Al and Fe from bauxite. Different parameters were evaluated in the SX step, such as extractant type, pH, temperature, and volume-to-volume ratio. The results indicated that using Alamine 336 as an extractant at pH 0 and room temperature led to an extraction of 90% of Fe and 15% of Al. Aluminum sulfate was completely precipitated as aluminum hydroxide and 55% of Fe using NH4OH. After the calcination procedure, obtaining a 99.9% pure alumina was possible without generating a red mud-similar residue.

Dec 28, 2023

By Lei Zhang, Zhibo Ma, Pingye Guo

The mechanical behavior of hard roofs often leads to significant deformation and strong mine seismicity in gob-side entry retaining (GER) of the longwall mining, affecting the mining safety and efficiency of coal mines. Field investigations indicate that the failure process of the retained entry by hard roofs can be divided into three stages. In this paper, theoretical analysis and numerical simulations are first used to evaluate the damage process of the retained entry under the hard roof. The results show that the instability of the retained entry in the conventional GER method results from the combined static-dynamic loads (in stage II and stage III) from the hard roof. Then, an innovative method, through fracturing roofs to maintain entry (FRME), based on GER with the hard roof is introduced to solve this problem. In this method, the roof of the mining area is pre-fractured, and when the coal seam is mined, the direct hard roof will collapse along the fractured surface due to its deadweight and mining pressure. The hard roof control effect on the static and dynamic loads, taking into account the innovative method, is verified with simulations and theoretical analyses. The results are compared with those of the conventional GER method. Furthermore, based on the numerical simulation study of dynamic responses, it is found that compared to the GER method, the dynamic response of the retained entry using the innovative method is more moderate and more conducive to the stability of the entry. Field application also validated the effectiveness of the innovative method to control the mechanical behavior of the “massive sudden collapse” of the hard roof.

Aug 14, 2023

By K. V. Lindell

THE 70-mile serpentine belt of eastern Quebec, producing 70 pct of the world's chrysotile, has 11 operating mines, two of which are underground, eight are open-pit, and one is both open-pit and underground. The combined operation, the Jeffrey mine, of the Canadian Johns-Manville Co., Ltd. is the largest producer in the group. Haulage All four of the separate operations falling under transportation at the Johns-Manville Jeffery Mine at Asbestos, P. Q., is handled by the company owned Asbestos and Danville Railway. A variety of equipment is operated over the 25 miles standard gage railway. Included in the scope of transportation are: 1-Hauling mill rock to the mill. 2-Hauling stripping and waste rock to the waste dump. 3-Hauling mill tailings to the tailings dump. 4-Hauling supplies from, and finished products to Danville. Main line tracks are permanent, but those within the pit and on the dumps have to be relocated constantly to follow the mining face and dump edges. Tracks within rock areas in the mine are laid on steel ties and can be moved by a bulldozer. Standard mechanical track shifters move tracks in dirt areas in the mine and on the dumps. Train-movements, both within the mine and over the out-lying tracks are controlled by a dispatcher located at the southwest rim of the pit. Steam locomotives haul all stripping, waste rock, and ore from the mine to the gathering yard at the northwest rim of the pit. Normally, two trains per operating shovel are used, with adjustment for optimum loading. In the event of shovel loading on the bottom level or in good stripping, service may be increased to three or four trains per shovel. From the gathering yard a 100 ton diesel-electric locomotive hauls the mill rock to the mill slips, a distance of about three miles. Electric or steam locomotives haul the waste rock and stripping to the waste dump located about two and one-half miles from the mine. Stripping cars are sprayed with a chloride solution and "gravelled" to prevent wet stripping from freezing in the cars during the cold months. Cars pass under a shaking screen which distributes an even

Jan 1, 1952

By S. Banisi, M. Pourkani, A. R. Laplante, A. Partoazar

"Canadian experience in industrial modeling of grinding, classification, and flotation kinetics was applied at the Sarcheshmeh copper mine, which processes 40,000 t/d of ore at 0.9 % Cu, in an effort to improve capacity and metallurgical performance. Mcivor's functional performance analysis was applied to ball mill cyclone circuits to maximize circuit capacity. Plitt's cyclone model was applied to predict how apex wear would affect classification performance, as part of a predictive maintenance program whereby the optimum wear period is determined for cast iron apexes. The flotation flowsheet was improved using Agar's optimization criteria. The impact of these technologies has been nothing short of spectacular, with an improved capacity of 1 S to 20""/o, and increased recoveries.1-IntroductionThe Sarcheshmeh copper ore body, which may rank as third or fourth largest in the world, contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and has been processing 40,000 t/d of ore since 1982. The mine produces 100,000 tonnes of copper and 2,200 tonnes of molybdenum concentrate per year. After three stages of crushing, the ore with a Fso of 12. 7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70""/o finer than 7 S microns. The concentrate of the first stage of flotation is reground and the tailings constitute the main component of the plant final tailings. A concentrate with an average grade of 31 % Cu is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83 and 87 % depending on the operating conditions and ore type.To improve the overall performance of the plant, it was first decided to evaluate the performance of grinding, classification and flotation. Measures were then taken to improve performance using the following criteria: (i) The improvements should be of the types that have resulted in significant improvements in other plants (e.g. Shannon et al, 1993); and (ii) The proposed improvements must be achievable with the manual operation presently used at Sarcheshmeh. The outcome of this study could also help in making decisions on the implementation of automatic control systems in the Sarcheshmeh copper mine."

Jan 1, 1998

By Russell Spry, Paul Schutz

The property of Pioneer Gold Mines of B.C. Limited is in the Lillooet mining district of British Columbia. It is on Cadwalleder Creek, 54 miles distant by motor road from Shalalth, a station on the Pacific Great Eastern Railway. Interests identified with the present company installed a small amalgamation plant on the property in 1924, consisting of a jaw crusher, Bryan mill and amalgamation plates, in which 75 per cent extraction was made on mill feed averaging 1.0 oz. gold per ton, the tailing being stacked for future treatment. That plant was operated until 1928, when the present company was formed and a 100-ton continuous counter current decantation cyanide plant was built. A second unit, having a nominal capacity of 200 tons daily, was added in 1932 and put into operation during September of that year. When handling ore comparatively free from slow-settling material, the combined capacity of the two units is 400 tons, the original plant now handling about 140 tons maximum. Production to date, from Pioneer, has been about $6,500,000, the present monthly production being better than $250,000 from about 370 tons ore daily, the recovery being 97 per cent. During the fiscal year ending March 31, 1934, a total of 112,936 tons ore and old amalgamation tailing was milled, the mill feed averaging 0.7670 oz. gold and the recovery amounting to 0.7422 oz. gold per ton, or 96.77 per cent. Milling costs, including refining and mint charges, during this period, amounted to $1.45 per ton of ore. Since the building of the first unit of the cyanide plant in 1928, direct cyanidation has been employed, no attempt being made to remove coarse gold from the circuit, other than that cleaned up from the grinding mills during relining and from the classifiers, as occasions permit. In the hope of still further increasing recovery, researches are going forward at this time with the view to removing some of the sulfides after primary grinding, the expectation being that selective treatment of this more refractory portion of the ore may hold economic possibilities.

Jan 1, 1935

By M Biesinger, R St C Smart

Diagnosis of the surface chemical factors playing a part in flotation separation of a value sulfide phase requires measurement of the hydrophobic and hydrophilic species that are statistically different between the concentrate and tail streams. Statistical methods, based on the monolayer-sensitive time of flight secondary ion mass spectrometry (ToF-SIMS) technique, have been developed towards this ultimate aim by measuring hydrophobic species (collector ions, dimers and metal complexes, polysulfides) as well as hydrophobic metal ions, precipitates and added depressant species. Reliable identification of specific mineral particles is central to this statistical analysis. A chalcopyrite/pyrite/ sphalerite mineral mixture conditioned at pH 9 for 20 minutes to study transfer of Cu from chalcopyrite via solution to the other two mineral surfaces, since this mechanism can be responsible for their inadvertent flotation in copper recovery, showed no statistical difference in the copper intensities on pyrite and sphalerite (selected from Fe and Zn images) after this conditioning. Principal component analysis (PCA) identifies combinations of factors strongly correlated (positively or negatively) in images or spectra from sets of data. In images, PCA selects these correlations from the mass spectra recorded at each of 256 ¦ 256 pixels in a selected area of particles. In the image mode, PCA has proved to be a much better method of selecting particles by mineral phase with clearer definition of particle boundaries due to multi-variable recognition. It has clearly separated a statistical difference in copper intensities between the sphalerite and pyrite phases. The PCA method has been applied to concentrate and tails samples collected from the Inco Matte Concentrator demonstrating extensive CuOH and NiOH transfer between the chalcocite (Cc) and heazlewoodite (Hz) minerals. Statistical differences illustrate the important discriminating depressant action of NiOH in flotation despite the activation of Hz by Cu transfer. The adsorption of the collector at specifically-identified Cu sites has been elucidated by the study. Importantly, the statistical analysis has been able to confirm some mechanisms and deny others proposed to control recovery and selectivity giving more focus on the control factors.

Jan 1, 2005

By A. J. M. Ross

BACKFILLING of stopes and other underground openings in the Homestake mine with sand tailings was undertaken primarily to reduce surface subsidence, which was wrecking much of the surface plant and a substantial part of the town, and to make possible maintenance of open-ings into the top of pillar stopes as mining progressed. The progress of this type of filling has been rapid and the results very satisfactory. Sand is now used in filling timbered stopes, shrinkage stopes, and abandoned drifts and crosscuts. A permanent pipe-line system for transporting sand tailings into the mine by gravity has been installed. The close filling of all stopes has resulted in important benefits other than the two major ones. The ore bodies in the Homestake mine occur as replacements in one distinctive layer of the metamorphosed pre-Cambrian sediments along a group of minor folds that together form a complex anticline that pitches at 35° to 40° to the southeast. The ore bodies are extraordinary for their magnitude. Widths of over 300 ft. have been mined and continuous ore has been followed for over 1200 ft. along the strike on some of the upper levels. Mining has been by shrinkage stoping with square-set mining of crowns and pillars. In the standard plan, shrinkage stopes 60 ft. in the direction of strike and the full width of the ore have been taken up to within 25 ft. of the level above. After filling, the crown has been mined by square-set methods, as have also the 40-ft. pillars between shrinkage stopes. In former days, when the supply of ore seemed inexhaustible, the mining of pillars was allowed to lag far behind because of the high cost. Many pillars were rather badly fractured on the level and in some places between levels. Coarse waste was used for backfilling all stopes, but in pillars it became increasingly difficult to maintain haulageways for the transporta-tion of this fill over the top of the stopes as mining progressed. The cost of maintaining these crosscuts was excessive and in many pillars prac-tically impossible. Therefore it became the practice to fill pillar stopes by tapping the back and sides of the stopes.

Jan 1, 1939

Cortez Gold Mines operated a conventional 2,100 mtpd (2, 300 stpd) cyanidation plant until the 5 million ton ore body was worked out in 1973. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME 1975.) Sub-grade material was dump leached at the mine -mill location in 1971, and the resulting pregnant solutions were added to the feed in the plant's Merrill-Crowe precipitation unit for gold recovery. In early 1973, however, Cortez put into operation their Gold Acres mine which was located about 13 krn (8 miles) to the west of the 'Cortez mine. Economics at that time dictated the heap leaching of the lower grade portion of the ore body with the higher grade being trucked to the Cortez plant for milling. Leach ore was moved directly from the open pits to prepared leach pads located on a gently sloping terrain of 570 grade. Preliminary pad preparation at the leach site included removal of vegetation which was composed of sagebrush and grading of the base. Pads were comprised of mostly minus 74 micron (200 mesh) materials and were made of clay-silt or slime tailings spread on the pad area in thin 50 to 75 mm (2 to 3 in. ) layers, compacted with a vibrating roller compactor to a depth of about 380 mm (15 in.). This was followed with a 100 mrn (4 in.) layer of coarse rock to prevent erosion on the pad and to allow solutions to move laterally under the heaps. Leach solutions normally traveled the length of the heaps and were collected in plastic-lined ditches from which they flowed by gravity to the pregnant solution storage pond. Both pregnant and barren ponds were constructed in the same manner as the leach pads - being lined with compacted tailings or clay silts. The Gold Acres facility required pregnant solution storage of 6 million liters (1.6 million gal) and barren storage of 3.8 million liters (1.0 million gal) designed to handle two to three days of heap drainage or to retain run-off from a major rainfall. Various heap heights were tried including 3, 6, and 9 meters (10, 20, and 30 ft), with a 6 m (20 ft) height selected. Ore was dumped by trucks in a single 6 m (20 ft) lift using a D8 Caterpillar dozer for

Jan 1, 1981

Discussion of the paper of Robert H. Richards, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 81, September, 1913, pp. 2299 to 2303. C. D. DEMOND, Anaconda, Mont. (communication to the Secretary*) :-Any one who has observed the increased efficiency of the Wilfley tables in a mill when fed from a reasonably good hydraulic classifier, as compared with tables fed from a spitzkasten., which is a very poor classifier, will agree to the advantage of ? better classification." It must be gratifying to Dr. Richards to know that, directly or indirectly, maim mill men are responding to his persistent advocacy of correct principles of ore dressing and are accepting the results of his investigations. As he says, these principles mean much for both the conservation of resources and the profit of operators. When Dr. Richards gives us accurate figures as to the maximum size of free mineral grains in the tailings of the Wilfley table and of the vanner, they will be of a good deal of practical value in delimiting the proper fields for the Wilfley, the vanner, and the round table. He has coined the terms ? free settling" and "hindered settling."' The series may be completed by adding "crowded settling," the main principle of action on both the vanner and the Wilfley table, which is the principle of separation in a mass of grains crowded as closely together as they can be and yet continue free to move between one another; and the occurrence of larger mineral particles in the Wilfley tailing than in that from the vanner is apparently caused by the more violent agitation, due chiefly to the purposely sudden reversal of the stroke. It seems likely that the agitation caused by the wash water dropping directly upon the concentrate on the vanner is more harmful than that due to the shake. For the very finest material that can be concentrated by usual methods, the round table is superior to the vanner because it does not, have even, the gentle agitation of the latter. The action on round tables is, not crowded settling," but "film sizing," 2 but agitation is harmful in the latter action. George Gates proved this harm 20 years ago : in slaking the first successful large-scale application of canvas tables in California gold mills he abandoned all shaking motion, although he had patented canvas tables to which such motion was applied.

Jan 11, 1913