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By C. F. Lucks, W. D. Wood, H. W. Deem

Electrical ad thermal-conductivity measurements have been made on titanium and 6 titanium alloys. The following expression was derived for calculating thermal conductivity from electrical conductivity: K = 0.626 s T x 10-8 + 0.00497 where K = thermal conductivity, cal sec-1 cm-2 cm0C-1 s = electrical conductivity, ohm-]cm-1 T = temperature, deg Kelvin This expression gives calculated thermal conductivities that are not in error by more than 10 pct from the experimental values for A-55 titanium and the 6 alloys measured; namely, A-110AT, C-110M, C-130AM, Ti-6A1-4V, Ti-140A, and Ti-155A, which are representative of the presently available commercial a and a-ß titanium alloys. THERMAL conduction in a gas, liquid, or amorphous solid is accomplished by energy diffusion from molecule to molecule, a relatively slow process. For this reason, such materials have a low thermal conductivity. When a material is crystalline, there is in addition a heat transfer by a vibratory motion of the crystal lattice. It is the contribution of this lattice conduction that makes a crystal a much better conductor than a gas, liquid, or amorphous solid. Lattice conduction, however, does not account for the relatively enormous thermal conductivity of a metal. This is explained by the freedom of movement of valence electrons. These valence electrons drift in a metal in a direction toward a drop in electrical potential and, to a lesser degree, in the direction of a temperature drop in the metal. Since both the electrical and thermal conductivities of a metal depend upon the presence of free electrons, there should be some relatively simple relation between the two. Wiedemann and Franz in 1653 first discovered an approximate empirical relation between the electrical and thermal conductivities of metals at the same temperature. This ratio varies appreciably with temperature, and Lorenz later showed that by adding a temperature factor to the Wiedemann-Franz ratio the value of K/(s T) is constant, where K is the thermal conductivity, a the electrical conductivity, and T the absolute temperature. This Lorenz constant shows a greater variation for alloys than it does for pure metals, the ratio increasing slightly as the proportion of alloying elements increases. It has been found that a closer correlation between electrical conductivity and thermal conductivity can be obtained by means of an equation of the form K - A s T + B [1] where A and B are constants. In this equation the term AsT is the contribution of electronic conduction and B is the contribution of lattice conduction to the total thermal conductivity. EXPERIMENTAL MEASUREMENTS Electrical-resistivity and thermal-conductivity measurements on titanium and 6 titanium alloys serve as the basis for determining an equation of the K = AsT + B type for titanium alloys. Specimens—Table I shows nominal compositions of the 7 metals on which electrical and thermal conductivities were measured. The electrical and thermal-conductivity specimens for each alloy were taken from stock of the same heat and usually from adjacent sections of the same bar. All of the metals measured were in a mill-annealed condition. Methods—Electrical-resistivity measurements

Jan 1, 1959

By H. Razavizadeh

The recovery of copper from oxide ores containing from 2% to 4.1% acid-soluble copper from the Sarcheshmeh mine in Kerman province of Iran was investigated by agitation leaching in sulfuric acid media. Quartz, as a major constituent, and malachite, as the major copper mineral, were identified in the ore by X-ray diffraction (XRD) and microscopic studies. A series of leaching experiments on the ore was performed over a concentration range of 10 to 20 g/L and temperatures from 15° to 50°C for a maximum of 3 hours. An initial high dissolution of copper was followed by a very slow reaction. By increasing acid concentration and temperature, the dissolution rate of copper was accelerated. However, by increasing acid concentration, the acid consumption of ores was also increased. Surface morphology of solid residues at various leaching times was examined with a scanning electron microscope (SEM). All SEM micrographs of the partially reacted particles displayed a needle structure. The kinetic studies showed that the malachite dissolution rate could be modeled in two stages. About 85% of the malachite dissolution occurred in Stage 1, while the remaining 15% occurred in Stage 2. The data obtained for the leaching kinetics indicate that the dissolution for Stage 1 was a diffusion-controlled reaction with an activation energy of about 20.6 kJ/mole, and the dissolution for the Stag 2 was a chemical-controlled reaction with an activation energy of about 87.6 kJ/mole.

Jan 1, 2008

By G. Hakes

A data bank on bolter technology was assembled which included state-of-the-art industrial equipment as well as new equipment currently under development for the U.S. Bureau of Mines. This information was augmented by field visits to selected USBM contractors, mine operators and others for the purpose of obtaining a thorough understanding of present coal mine roof bolting safety and productivity problems. These direct contacts provided valuable insight regarding requirements for new bolting equipment necessary to assure wide user acceptance. Using this background material and information, a series of roof bolter concepts was developed offering improved operator safety with minimum impact on productivity. The concepts developed utilize existing componentry and technology wherever possible. All of the concepts feature a fully protected operator station located 6' to 10' outby the active bolt line and the hazardous rotating machinery.

Jan 1, 1978

By Bradford F. Townsend, Matthew E. Fowler, Michael F. Smithson, Lawrence R. Eckert

All Rights Reserved. Printed in the United States of America. Information contained in this work has been obtained by SME, Inc. from sources believed to be reliable. However, neither SME nor its authors guarantee the accuracy or completeness of any information published herein, and neither SME nor its authors shall be responsible for any errors, omissions, or damages arising out of use of this information. This work is published with the understanding that SME and its authors are supplying information but are not attempting to render engineering or other professional services. If such services are required, the assistance of an appropriate professional should be sought. No part of this publication may be reproduced, stored in a retrieval system, or transmitted in any form or by any means, electronic, mechanical, photocopying, recording, or otherwise, without the prior written permission of the publisher. Any statement or views presented here are those of the authors and are not necessarily those of SME. The mention of trade names for commercial products does not imply the approval or endorsement of SME.

Dec 1, 2010

This paper deals chiefly with the preliminary examination, when a decision has to be made as to whether a prospect has a chance of developing into a mine or not. At this stage it is important to have a fairly clear conception of what is expected of the prospect, andjust what type of mine is sought. Much money can be wasted testing a prospect suitable for a prospector or small syndicate, but which never looked like becoming a mine of sufficient size to interest a company. When undertaking an inspection, therefore, it is necessary to have a broad outline, at least, of the policy of the principals for whom the report is intended.At this early stage it is frequently necessary to form an opinion of the prospects of a property on very meagre data. When new to this type of work one's first reaction is often th~t it is impossible to pass an opinion on the data available. It. is absolutely necessary,however, that an opinion of the possibilities and chances of favourable development be given; The report should strive to convey this information, making quite ::lear the risk involved, how complete are the data on which this opinion is based, and the possible magnitude of the prize. A Preliminary report has to state whether a property has sufficient prospects to warrant testing.The following procedure is suggested in order to make best use of the data and minimise hasty conclusions influenced by preconceived ideas-an ever-present danger.Examination.The first stage is the examination. This should aim at collecting all available significant data, but avoid clouding the issue with detail which, at this stage, will not affect the decision. The following points are considered to be the main ones for which to look:-Geological Environment.As good a working knowledge as possible'of the geology of the area should be obtained. Government Geological Survey reports and maps, if available, will be a big help. A detailed knowledge of the regional geology is not necessary a,t this stage, but a broad outline of the geological setting of the prospect under review, and also of any other important mines in the district, is often helpful in the final decision.The Nature of the Ore-Body.Any facts concerning the nature or occurrence of the lode should be noted as, coupled with experience of similar bodies, they affect the ultimate decision.Structural Control.This is probably the most important factor, in the light of present geological knowledge. Its chief importance is due to two facts:-(a) Structure plays one of the prime parts in ore deposition.(b) Geological knowledge, of structure is somewhat in advance of knowledge of other factors influencing ore deposition.

Jan 1, 1950

By Warren L. J

The effectiveness of shear-flocculation as a method for upgrading simple mixtures of ultrafine (1-11 \lm) scheelite and garnet was evaluated. Statistical methods were used in the design and analysis of the main test program.The stirrer speed, stirring time, and initial weight ratio were all found to have significant effects on both the recovery and grade of the floes produced. The particle sizes also had an effect on the final grade. There were several important interactions between factors which also affected the results. Optimum conditions for shear-flocculation were established, and under these conditions about 80 per cent of the scheelite could be recovered at a grade which depended on the composition of the starting mixture. Starting with mixtures containing 10 and 50 per cent scheelite, the best final grades were between 20 and 50 per cent and between 70 and 90 per cent respectively, the actual value depending on the particular combination of particle sizes.

Jan 1, 1977

Detailed characteristics of gold from Pliocene auriferous quartz gravels from Waimumu (Southland) have been studied using a combination of optical microscopy, energy dispersive microprobe analysis and scanning electron microscopy techniques. The gold at Waimumu is classified as either: 1. detrital gold, or 2. irregular gold. A number of different morphologies a recognisable within each 'gold type'. Evidence is given for an predominently authigenic origin for some of the gold at Waimumu.

Jan 1, 1987

Three complete connector assemblies and ten separate contacts were submitted to a series of tests to establish the connector permissibility and compliance with the Bureau of Mines' research Contract H 0110423. Test Procedure TPC-0100 was coordinated, and the test plan approved by the Bureau as fully complying with the intent of the contract for the low voltage, medium voltage and high voltage A.C. connectors. The components and connector assemblies have been demonstrated in testing that they meet or exceed all the requirements of the test procedure.

Jan 1, 1972

This Bureau of Mines publication contains five papers pertaining to the effective grounding of mine power systems, as follows: Ground-check monitor types and safety characteristics; ground wire monitors for surface coal mines; ground bed design and measurement; electrical ground rod corrosion; and a continuous ground bed resistance monitor.

Jan 1, 1979

By L. T. Stanley

A method is described whereby the behavior of the front is approximated when gas is injected into a thick reservoir having reasonably homogeneous properties. The method is applied to the Abqaiq field of Saudi Arabia. The producing member is the Arab-D, a high-permeability calcarenite which has no continuous betlding planes. Field performance indicates the injected gas to be partially miscible with the reservoir oil because gas volume factors experienced are lower than those predicted by the gas law. Eqlrilibril~m calculations confirnz ihis and also show that the volume factor may vary appreciably, depending on the relative quantities of oil and gas in contact in the reservoir; consequently, volutnetric balances are made for the purpose of determining the behavior of the gas volume factor during injection history. The effect of gravity segregation on position and shape of the gas front is approximated by dividing the producing section into three zones vertically and tracing the movement of the average interface in each zone, with counterflow of oil and gas taking place betweell zones behind the front. This results in a gas front that advances more rapidly along the top of [he formation than along the base, producing an "umbrella" effect in cross section that becomes more pronounced as the front progresses. The calculations cover the gas-injection period of Abqaiq field history through 1958, and the position of the gas front is plotted at one-year intervals. Finally, some comparisotls between the calcrlated fronial position and the gas-oil contact as Measured by neutron logs are made. INTRODUCTION Fig. 1 is a structure contour map of the Abqaiq field. For analytical purpcses, the field has been divided into two areas designated A and B, as shown on the map. Gas injection into Area A was begun in early 1954 into two wells located near the crest of the structure on the north-south axis of the anticline. During the period covered by this paper, the average injection rate has been 122 MMscf/D with a maximum sustained rate of 160 MMscf/D. The reservoir pressure has been maintained at 2,450 psig since start of gas injection. The reservoir originally was undersaturated but, prior to pressure maintenance, the pressure was drawn below the bubble point in most of the Area A section, creating an in-place free-gas saturation of 8 per cent at the crest of the structure. The average critical gas saturation is 13 per cent established by laboratcry core tests. That the entire reservoir remained below critical gas saturation prior to gas injection is borne out by the fact that none of the crestal wells have exhibited high gas-oil ratios during production history. The Area A reservoir is a carbonate section about 200-ft thick having average porosity and permeability of the order of 22 per cent and 500 md, respectively. The productive horizon is the Arab-D member of Jurassic age and is encountered at an average depth of

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By Diane Jones, Jeremy Spearman, Tom Matthewson

Requirements for minerals are on the increase globally and the deep sea offers a potential source of essential materials. Extracting these minerals from what is considered an environmentally sensitive area, and one about which little is still known, is a challenging activity, which needs careful consideration and an appropriate environmental impact assessment to be carried out. Extraction of material from the deep sea beds can cause disturbance or damage to benthic communities; impact on mobile marine species such as fish, mammals and reptiles; cause increases in suspended sediment and sediment deposition onto potentially sensitive areas or potentially release contaminants which were previously bound to the sediment. This paper discusses the potential environmental issues surrounding extraction of deep-sea deposits using aggregate dredging techniques and how those issues can be appropriately assessed. In particular issues of suspended sediments, plume transport, water and sediment quality and underwater noise impacts on potentially sensitive species and habitats will be explored.

Sep 1, 2014

By Walker G. S

A study of the crystalline materials on or near the surface of freshly made or "green" superphosphate granules and mature, "set" superphosphate granules was undertaken using X-ray diffraction, X-ray mapping, scanning electron microscopy and electron probe micro analysis techniques. Calcium sulphate (CaSO4), calcium sulphate hemihydrate (CaSO4 1/2H20), and mono calcium phosphate monohydrate (Ca(H2PO4)2H2O), were the only crystalline species present in the "green" superphosphate.In the set superphosphate however, additional compounds were present, namely complex iron and aluminium phosphates Ca(Fe,AI) H(HPO4hF2)2H2O, (Fe)3(H3O)H8 (PO4)6H2O and calcium sulphate dihydrate (CaSO4.2H2O).The complex phosphates tied up the major part of the Fe and Al in the system and were concentrated on or near the surface of the granule. An intimate mixture of CaSO4.2H2O, and CaSO4, along with the above mentioned complex Fe-AI compounds, crystallize close to the surface of the granule and cause the setting of superphosphate that takes place during storage.

Jan 1, 1980

By Sunil Koppalkar, A. R. Laplante

"A Pilot Plant was installed in one of the grinding circuits of the Dome Mill, Ontario, Canada. Tests were conducted on a KCCD12 Knelson Concentrator to systematically analyze the effect of operating variables on the performance. The main objective of this test work was to explore the effect of very coarse (6 mm) feed size on GRG recovery and performance of the Knelson concentrator. Preliminary results have shown that coarse feed size is deleterious to maximizing recovery. Size-by-size GRG recoveries have shown an unusual trend in the sense that recoveries do not decrease monotonically with decreasing gold particle size. Similar trend observed in one other full-scale unit has been compared. The observed trend is probably linked to the feed size distribution and top size.INTRODUCTIONKnelson Concentrators are the most widely used semi-continues centrifuge gravity separation units for the recovery of gold and platinum group minerals (Xiao and Laplante 2004). The main advantages over the competing centrifugal gravity units are their design, operating simplicity and good separation efficiency that make them the most common choice for gravity circuit. Despite some characterization at bench scale and a limited characterization at full scale, there is a lack of a comprehensive study of impact of operating variables on efficiency that reflects realistic full scale operating practice. GRG characterization studies, at bench scale, on these units provide information about the occurrence of gold in ore samples but fail to provide adequate information about the effect of operating variables for full-scale units. It is well known fact that such work is difficult to undertake on full-scale units owing to the difficulties involved in sampling and systematic variation of operating parameters. To tackle these problems, a pilot facility consisting of a 12-in CD Knelson Concentrator, a feed screen and tailing pump was installed in one of the grinding circuits of Dome Mill. This paper describes the work carried out and the results.Porcupine Joint Venture (PJV)Porcupine Joint Venture was formed in 2002 between Placer Dome Inc. and Kinross Gold Corporation. The JV utilizes the milling capacity of the Dome Mill to process ore from Placer Dome and Kinross mines. The Dome Mill circuit has gravity concentration, cyanidation and CIP circuits to recover gold. Gravity recovery has been used at Dome Mill since inception to recover gold: first by using conventional jigs, which were eventually replaced by two Knelson Concentrators in 1993, one in each of the grinding circuits. Three more Knelsons joined the circuit in 2002 to improve gravity recovery. An Acacia Intensive Leach Reactor was also added in 2002 to process gold concentrates replacing shaking tables (Chong et al 2004)."

Jan 1, 2007

By Thomas M. Barczak

The resultant load vector is the representation of the forces applied to a longwall roof support element by caving strata into a single, quantifiable measure of support resistance. The relatively complex kinematics of the shield support prohibit a determination of support resistance simply by summation of leg forces. A method is being investigated by the Bureau of Nines to determine the resultant load vector by instrumenting supports with pressure transducers and strain gages to measure leg, canopy capsule, and lemnisoate link forces. This concept has been laboratory tested in the Bureau's Nine Roof Simulator. Functional relationships among variables have been assessed, and confidence intervals have been established for prediction of the resultant load vector parameters.

Jan 1, 1984

By D. R. Dolinar

Geogrids can be used as surface control on both the roof and rib in underground mines to prevent the fall of small pieces of rock and material from between the primary supports. The use of screen material such as geogrid can significantly reduce the number of ground fall injuries caused by these small rock falls. Because of the potential for reducing injuries through the use of screen, the National Institute for Occupational Safety and Health (NIOSH) has conducted tests on screen materials including geogrid to determine their support characteristics and identify parameters that can affect their performance with respect to maintaining broken or failed rock. A specially designed test frame installed in the NIOSH Mine Roof Simulator (MRS) has been used to evaluate the load and stiffness characteristics of the screen material. Two types of geogrids with various designs have been tested-a polyester yarn and a polypropylene grid. The load-displacement characteristics of these grids were determined and the load capacity and stiffness were calculated through 20in of displacement. The load capacity and stiffness of the geogrid in general are shown to be related to the ultimate tensile strength of the screen material.

Feb 27, 2013

By P M. Ashley, M Muller

Massive sulphide ZnPbCuAg deposits in the Halls Peak district of north-eastern NSW are of the submarine volcanogenic type. They occur as a number of small lenses enclosed in shale and epiclastic rocks, in association with felsic volcanics of early Permian age. The deposits crop out in steep terrain near the eastern edge of the New England plateau. Local relief of 700 m and slopes up to 50¦, coupled with an average annual rainfall of about 1000 mm has led to local severe land sliding in previously mined areas. Mining from the 1890s to the 1960s has also resulted in unconstrained drainage of heavy metal-bearing acid waters and physical dispersion of finer grained metalliferous sediments into the Chandler River. Waters emanating from the main landslide area, from underground workings and draining scattered unconfined stockpiles of massive sulphide ore have pH values of 3 to 4 and contain elevated contents of Zn, Cu, Pb, Cd and Fe. Chemical mobility of heavy metals into streams occurs under low pH conditions with Zn exhibiting the greatest dispersion. Increasing pH conditions downstream cause precipitation of Fe phases and coprecipitation of heavy metals. Physical dispersion of secondary metal-bearing minerals into the surrounding creeks results in wt% levels of Cu, Pb and Zn in the stream sediments thereby exceeding the regional background values by one to three orders of magnitude. Areas disturbed by mining are characterised by lack of vegetation, biomagnification of heavy metals into algae and lower plants, and partial recolonisation by metal-tolerant grass species (Snow Grass, Poa sieberana; Blady Grass, Lomandra longifolia). Rehabilitation measures will have to include landslide stabilisation, impervious sealing or removal of sulphidic materials, import of topsoil and revegetation by metal-tolerant, slow-growing native grass and shrub species.

Jan 1, 1995

By F. Calizaya

This paper describes the basic requirements for the utilization of belt air to ventilate working sections at remote locations in relation to the main fans. It highlights the risks involved, potential problems, and the provisions that must be followed by mine operators for the safe use of belt air. The paper also presents a brief review of the current regulations, conveyor belting tests, atmospheric monitoring system requirements, and the experience gained by mine operators to utilize the belt entries to ventilate remote working sections or areas where mechanized equipment is being installed or removed. The paper concludes with a set of recommendations with respect to the utilization of belt air and the composition of belt material in under-ground coal mines.

Jan 1, 2008

The value of domestic raw nonfuel mineral output in 1985 was estimated at $23.7 billion, 2% above the 1984 value of $23.1 billion. This crude mineral output was the basis for approximately $244 billion worth of processed materials of mineral origin. [ ] A U.S. company permanently closed its alumina plant in St. Croix, U.S. Virgin Islands, reducing domestic alumina capacity by 10%. Another U.S. company sold its 26% interest in an alumina plant in Jamaica to its two U.S. partners. On completion of the purchase agreement, the two remaining partners promptly closed the plant for an indefinite period, citing a world oversupply of alumina and unacceptably low prices.

Jan 1, 1986

By Mike Sapiro

"Heath Steele Mines is a base metal producer situated near Newcastle, N.B. presently treating 3500 TPD of complex Cu-Pb-Zn ore. The ore consists of massive fine grained sulfides averaging I. 1% Cu, 1.6% Pb and 4.5% Zn.The original mill started in 1956 at 750 TPD and was successively expanded through 1500 to 3000 TPD in 1969. In 1972 an expansion to 4000 TPD was initiated. It was planned to obtain the maximum benefit from equipment and manpower by adding some process equipment, instrumentation, an on stream analyser, and a process control system. The benefits from this approach could be greater than normally expected, because of the complex metallurgy of this ore.IntroductionThe crushing plant is of conventional design with two 5 ½ ft, Symons cone crushers with the crushed ore screened to -7/8 inch on a 61 x 16 1 Al I is Chalmers vibrating screen and stored in four 1000 Ion bins. There are two para I lei grinding circuits, each one consisting of the following: one Rod Mill in open circuit, one Ball Mill in closed circuit with two 20"" Krebs cyclones, and one Ball Mill in closed circuit with three 15"" Krebs cyclones, Al I the Mills are 9' x 121 Al I is Chalmers Mills. The secondary cyclone overflow from each grinding circuit is thickened and combined to form flotation feed. The final product is 65% - 400 Mesh. Two regrind mills have been installed as part of the expansion."

Jan 1, 1977