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By L. A. Pursell

The case history, which is the subject of this paper, involves the expansion of a lime plant in Texas. The major new production unit is to be a 545.45 tonne per day coal fired rotary kiln. Auxiliary facilities include additional quarry haulage equipment, modifications to the crushing plant, a hydrator, storage facilities, and a loading system. The existing plant consists of a high calcium limestone quarry, two coal fired rotary and 10 gas fired shaft kilns, nth crushing, screening, stockpiling, hydrating, storing and loading facilities. The proposed modification would increase quarry volume from 545,000 tonnes per year to 805,000 tonnes per year. The plan also includes the addition of dolomitic lime to the product line which requires the purchase and handling of 186,000 tonnes of screened, washed dolomite yearly. A decision to undertake such an expansion was previously a simple and straight forward process. Such a decision now faces a major and complicating obstacle which can delay the project, raise the expected cost, or prevent it entirely.

Jan 1, 1980

By T. W. West

A description of a coal washing spiral is presented, together with a discussion of the parameters which define the envelope within which the spiral can be successfully applied. The correct process engineering, installation, operating and maintenance parameters are addressed as well as actual results for new and retrofit installations on USA coal capital and operating costs of spiral circuits are compared to those of traditional fine coal circuitry.

Jan 1, 1988

By R. H. Smith

The total energy demand in the United States increased about 4 per cent annually during the 1960's 1 and is expected to increase at the same rate during this decade. Z Fuel oil growth in the 1960's was only Z per cent per year until 1968, but it increased to 8 per cent in 1969 and 11 per cent in 1970. Fuel oil demand is expected to grow at a 6 per cent rate through the 1970's. Part of this rapid increase in demand has been caused by restrictions placed on the sulfur content of fuels. High sulfur coal can no longer be burned in many parts of the country and has been replaced by low sulfur fuel oil. An insufficient supply of natural gas has also contributed to fuel oil demand. Air pollution regulations have not only caused an increase in fuel oil consumption, but they have also caused a reduction of sulfur content in residual fuel oils. Current regulations call for 1 per cent maximum sulfur content fuels over much of the country. Sulfur contents of fuels are limited to 0. 3 per cent in certain areas on the east and west coasts. Sulfur in residual fuel oils from high sulfur imported crude oil is reduced to 1 per cent currently. However, sulfur reduction to 0. 3 per cent is difficult and expensive.4 As a result, refiners may be hard pressed to supply the market for low sulfur fuel oil. The oil shale deposits in Colorado and Wyoming potentially are a huge source of low sulfur fuel oil. It will be shown that the low sulfur content and more importantly the sulfur

Jan 1, 1972

By Santhanam Sundaram

The potential for the utilization solar energy for the production of industrial process steer in mineral processing industries is discussed briefly. To advance the technical and economic feasibility of solar-thermal systems applications, U.S. Department of Energy has funded several projects. NL industries of Newberry Springs, California is one among them and has been chosen as a case study. The NL Industries solar system consists of 950 square meters of Del, single-axis, tracking, parabolic-trough, concentrating collectors. It was designee to produce a portion of the 420°. 530°k. steam utilized in a trying operation reduce the moisture content of hectorite ore from 10 percent to 4 percent. (Hectorite is an hydrous magnesium silicate which, when refined, is of significant commercial interest because of its applications in various chemical and food processes.) is estimated that implementation or this solar system could result an annual savings of 3.545 billion KJ ,3.360 billion BTUs), or the equivalent of 90 cubic meters (600 barrels) of oil and a net reduction of 1,860 kilograms (4100 pounds) of air pollutants annually. The technical, economic and institutional issues encountered in the course of this project are also discussed. Finally, the impact of the commercialization of solar energy applications in mineral processing industries is evaluated.

Jan 1, 1981

By Oscar M. Wicken

Iran was a lodestone for the attention of industrial countries and companies wanting to share in its industrial development during the last two decades and particularly after 1972. Prior to 1972, most industrial projects in the private sector were on a relatively small scale and could be carried out in reasonably orderly fashion in spite of limitations of a primitive infrastructure and meager indigenous supply of managerial and technical skills. People with these skills could be found but not in abundance, and many of them were total¬ly engaged in various government ministries and projects. Thus it was that nearly all private sector industries, even long-established ones, required one or more foreigners to give guidance. A barite mine and processing facility was brought on line by Dresser Industries in 1960 without serious operational problems. A few expatriate experts were able to select and train a stable group of staff and production people and establish sound management principles for continued growth. The huge and increasing oil revenues after 1973 gave superficial credence to ideas of the Iranian government that the country could easily finance the jump into the industrial 20th century. All necessary skills and equipment not available in Iran would be imported, and intensive training programs would prepare Iranians for operating the new facilities. Large projects for steel production, copper production, oil refining, petrochemical production, cement production, and others were started or expanded. These are basic industries that consume substantial quantities of refractories. Studies made in 1972 indicated that the existing industries plus expanded and new projects scheduled to come into production would lead to a market sufficiently large to justify a refractory producing plant. Concurrent studies of raw material sources showed that clay, the largest tonnage item, would be locally available; the other raw materials would be imported. Iran Refractories Company (IREFCO) was founded in 1973 and started production in 1976.

Jan 1, 1981

By John F. Martonik

When OSHA came into existence in 1970, it adopted national consensus standards to improve safety and health conditions in work places. Coke oven emissions, which are proven to cause cancer and respiratory diseases, were regulated by the coal tar pitch volatile standard, adopted from the 1969 edition of the Threshold Limit Values for Airborne Contaminants by the American Conference of Governmental Industrial Hygienists. The coal tar pitch standard, based upon the benzene soluble fraction of total particulates, limits the personal exposure to 200 micrograms of the benzene soluble fractions of total particulate per cubic meter of air. This standard can be classified as a performance standard because only general limits are set which allows the employer to choose for themselves the means to comply. Both the United Steelworkers of America and the American Iron and Steel Institute petitioned the Secretary of Labor to write a new standard for coke oven missions, each thinking the coal tar pitch standard unsatisfactory.

Jan 1, 1980

By A. English, Edward Martinez, L. A. Haas, N. W. Johnson, Harrison Copper, M. E. Tawil, C. C. Harris, G. V. Jergensen, D. Halbe, M. P. Amsden

Declining US ore grades along with skyrocketing energy, materials, labor, and regulatory costs underscored the urgency for technological innovation in the mineral processing area. Re¬liance on the scale-up and automation of unit operations was the sector's primary battle plan. Debate continues on the upper size limit of concentrating plants and whether process controls should be centralized or distributed in the mill. In the materials handling area, the economics of large-scale and mobile systems, particularly conveyors, were well documented in the literature. Comminution, the most costly and energy consuming area of the processing flowsheet, saw improvements in process controls and liner designs. The trend toward larger grinding units stalled somewhat during the year owing to some operational problems at commercial plants. Research and development of apparently less expensive physical separation processes-that also boast recovery of cleaner and more environmentally acceptable products -gained momentum during the year. Because of escalating capital cost requirements for new plant capacity, the year saw a greater interest by mill designers to re-examine existing capacity. The focus here is clearly to modify or retrofit unit operations with more efficient components as an alternative to installing more costly new equipment. The problems facing the processing sector during the 1970s can be expected to continue in the decade ahead. Technological gains from scale-up and process controls appear

Jan 5, 1980

By J. R. Sturgul

A dynamic input-output model for a coal power enterprise that now includes four underground coal mines, three coal washing plants, a power plant and a railway transportation system was developed. The model is applied to analyze investment alternatives for long-term development, including the expansion of existing mines, the opening of new mines, the construction of new coal washing and power plants and the upgrading of the railway system. As a result, the optimal investment strategy with capital schedules and projected sector outputs is obtained. Also, the importance of the new power plant to the strategic development of the company is demonstrated. It is concluded that, without applying a dynamic input-output model, it would be much more difficult or impossible to conduct such a systematic and compete analysis.

Jan 1, 1999

By Harold Bloom

The Mount Pleasant Mines, or Mount Pleasant Joint Venture as it is presently known, is located approximately 59.5 km (37 miles) south of Fredericton, NB, Canada. Earliest indications of mineralization were found in 1954, during a regional reconnaissance geochemical stream sediment survey carried out over large parts of New Brunswick and the Gaspé Peninsula. The regional geology of New Brunswick is shown in Fig. 1. This program, sponsored by Selco Exploration Co. Ltd. of Toronto was conducted by H. Bloom, H.E. Hawkes, J.E. Riddell, and J.S. Webb. At its conclusion in 1955, 30 significant anomalous drainages were identified for follow-up evaluation. A total of 4937 sample sites were occupied. Geochemical reconnaissance using stream sediment sampling was a new technique which had only been field tested the previous year in the Gaspé and the northern coastal area of New Brunswick. Sediments collected from known mineralized drainages were found to leave a train of anomalous heavy metals from 1.6 to 9.7 km (1 to 6 miles) in length. When the exercise was extended to a reconnaissance survey, the previously unknown mineralized prospect at Nash Creek, New Brunswick, was discovered, staked, and later drilled (Hawkes and Bloom, 1956). Although it did not come into production, the discovery left little doubt as to the viability of the method and led to the Selco sponsorship of the program for the following two years.

Jan 1, 1991

The concept of "classification is expanded beyond the performance of the classifying equipment to include all aspects of the fines removal system of the grinding circuit. Design and operating parameters which affect fines removal are identified and discussed, and a quantitative index for circuit fines removal effectiveness is defined. This approach lends new interpretation to certain observed phenomena, and suggests useful guidelines for circuit design and operation.

Jan 1, 1985

By Christopher J. Bise

Traditionally, mine drainage systems have not received the high level of emphasis on careful engineering design which has been rendered to many of the other auxillary operations. This may be due, in part, to the situation-response approach toward dealing with an influx of water. When the installed system becomes inadequate, costly alterations to the pump, such as staging or speed increases, are considered expedient solutions. In order to rectify this situation, computer-aided solutions to mine drainage problems could represent an efficient alternative to the methods presently used. This paper will consider the application of microcomputer technology to mine drainage problems and will show how the relative economy and convenience of microcomputers can make them become valuable analytic tools at any mine site.

Jan 1, 1983

By Boris J. Kochanowsky

In the mobile crushing unit system trucks are replaced by a conveyor resulting in a "continuous" almost "fully-automated" operation with all its attendant advantages. In consequence, overall operating costs are greatly reduced. Because of the new designs of large mobile crushing units that have been brought on-to the market, the use of these plants has increased significantly during the last three years. This paper describes the designs of new mobile crushing units, and emphasizes the advantages that such plants offer to the rock products and mining industries. It examines operating costs, and indicates the conditions in which the application of these plants is most profitable.

Jan 1, 1968

By D. Carr, C. Anderson, M. Deal, C. Larson

"Construction of the OceanaGold Haile operation’s process plant in Kershaw, South Carolina is currently under way, with ore through the mill expected in late 2016 and commercial production expected in early 2017. The process is designed to treat 6,350 tonnes per day with minor upgrades required to achieve 8,270 tonnes per day. Plant optimization during detailed engineering, combined with a focus on de-risking the process, has resulted in numerous process changes since the initial design. This paper describes the plant optimization and de-risking, it also presents a proven commissioning strategy modelled after OceanaGold’s Didipio operation, whilst identifying potential expansion opportunities. INTRODUCTION Haile Gold Mine is located in Kershaw, South Carolina, approximately 50 miles south of Charlotte, North Carolina. Gold was first discovered at the site in 1827, and the property has been through several operating periods since. The last operation employed heap leaching from 1985 through 1992. Since then, the property has been closed and fully reclaimed. Romarco Minerals acquired the property in 2007, and began an exploration program that delineated 2.02Moz. of gold in reserve at an overall grade of 2.06 g/t. OceanaGold acquired Romarco Minerals and the Haile Operation in October of 2015. OceanaGold operates under a risk management philosophy in all aspects of operation, both corporately and at each individual business unit, and has implemented numerous process changes to eliminate or mitigate risks identified in the initial design of the processing circuit, and provided increased opportunity for optimization and expansion. This paper describes the plant optimization and de-risking, it also presents a proven commissioning strategy modelled after OceanaGold’s Didipio Operation, and identifies potential expansion opportunities. INITIAL PROCESSING CIRCUIT DESIGN An extensive metallurgical test work program was undertaken by Romarco Minerals and third party laboratories during the initial design phase of the Haile project [1]. This preliminary test work provided the technical information as a basis of plant design. The initial flow sheet for the Haile processing circuit is shown in Figure 1 (see APPENDIX). The plant is designed to process 6,350 tpd (7,000 stpd) on an annualized basis. The process begins with a jaw crusher followed by a SAG and ball mill grinding circuit to reduce particle size to P80 74 µm. A flash flotation cell treats 50% of the cyclone underflow in the ball mill circuit. Grinding circuit product is sent to a rougher flotation circuit to recover gold-bearing pyrite. The flotation concentrates from flash and rougher flotation are reground to P80 13 µm using stirred media detritors (SMD’s). The rougher flotation tail is pumped to the carbon-in-leach (CIL) circuit for gold recovery. The reground concentrate is pre-aerated prior to being pumped to the CIL circuit. The CIL circuit is a hybrid circuit with eight tanks. Tank 1 leaches the pre-aerated concentrate only, and tanks 2 through 8 leach the combined flotation tailing and tank 1 discharge. The carbon strip circuit is a standard pressure Zadra system, followed by electrowinning, and refining using an induction furnace. CIL tailing passes through a cyanide recovery thickener prior cyanide destruction and then pumped to the tailing storage facility (TSF)."

Jan 1, 2017

By J. Guzman

Introduction The Carlin Trend is North America's premier gold producing district. It is located in northeastern Nevada's Elko and Eureka Counties along a northwest trending belt about 65 km (40 miles) long and 8 km (5 miles) wide (Thorstad, 1989; Jones, 1989). This trend is the worldwide reference site for epithermal, sedimentary rock-hosted microscopic gold deposits. At least 19 deposits have been discovered to date, varying in size from 933 kg to 1.08 kt (30,000 to 35 million contained oz) of gold (Fig. 1). Newmont Gold Co. and its parent, Newmont Mining Corp., jointly constitute the largest mineral right holders in the district. They own or control more than 1000 km' (386 sq miles) in and around the Carlin Trend and own all or part of I6out of the 19 mines and prospects identified to date. Since the initiation of Newmont's exploration activities in the Carlin Trend in 1961, 2.24 kt (72 million oz) of cumulative gold resources have been identified. Cumulative production from all mines since the start-up of Newmont's Carlin mine in 1965 to the end of 1989 was about 202 t (6.5 million oz) (Jones, 1989). The incentive of sustained high gold prices and innovation in processing technology resulted in a significant acceleration of gold output over the last few years. Newmont Gold alone produced more than 43.5 t (1.4 million oz) in 1989. That is equal to 22% of the cumulative 1965 to 1988 output, and an almost 200% increase over its 1986 output. The same incentives produced even more spectacular exploration results. In each of the last five years, net additions to reserves and resources outpaced current production by substantial margins. These facts demonstrate the spectacular past prospectiveness of the Carlin Trend and the success of focused, multi-disciplinary exploration methods that made it possible to more than offset the recent accelerated depletion of gold resources. However, is this situation sustainable? How long can the mining companies along the Carlin Trend keep on finding resources faster than they deplete them? These are some of the questions that motivated this study. The authors have not quantified the future potential for gold exploration in the Carlin Trend nor established a deposit discovery path. But strong indications were discovered that the [ ] Carlin Trend remains a relatively immature exploration district and that the potential for significant new discoveries is high. Methodology and data The approach chosen to address the above questions was simple. The authors compiled deposit size data, measured in contained ounces of gold resources, for all known deposits along the Carlin Trend (Table 1). The resource information was obtained by adding cumulative historical production (adjusted for mining losses and metallurgical recovery) to 1989 year-end published resource inventories. In a mature exploration area, where most deposits have been discovered, this distribution would be expected to approximate lognormality and would plot along a straight line on a lognormal probability scale. This result was found in previous work by Allais (1957) and recently confirmed by Cox and Singer (USGS, 1986) in regard to various types of mineral deposits in several regions of the world. It was also found to hold true for oil and gas pool size distributions (Arps and Roberts, 1958; Kaufman, 1962; McCrossan, 1969). [ ] The data used were compiled by Newmont Exploration geologists. The purpose of the study is to make inferences about the underlying geologic processes in the district and the maturity of the exploration effort. Therefore, deposits were not classified according to ownership but according to geologic occurrence as known from current information. Newmont's Post and Barrick's Goldstrike and Betze deposits, for example, are shown as a single occurrence to reflect the actual geologic setting. The cumulative frequency distribution of deposit sizes was plotted on lognormal probability paper (Fig. 2). The abscissa shows the cumulative fraction of deposits at or below a certain deposit size and the ordinate shows the deposit size in thousands of ounces of contained gold resources.

Jan 1, 1992

By Thomas E. Howard

Engineers today face a tremendous challenge. Growing population and rising standards of living all over the world are creating unprecedented and accelerating demands on and for all kinds of technology. At the same time, the public is becoming more and more conscious of--and concerned with--the impact of technology on human welfare. The growing controversy over the Supersonic Transport is a case in point, It is clear that the modern engineer must expect to spend a growing share of his time considering this dilemma and evaluating the tools that are available to him, not only with respect to their adequacy for the job at hand but also their capability for meeting the increasingly demanding--and, in many cases, conflicting--requirements of the future. This meeting seems a particularly appropriate place to speculate about the future of mining technology. The summary of the state-of-the-art in surface mining contained in the papers scheduled for the International Conference, together with the recently published Surface Mining volume, constitute not only a major contribution to the technical literature but also form an impressive record of technologic progress. That record, and the advanced technology displayed in the exhibits, and to be discussed in the SME technical sessions demonstrate engineering achievement in which the mining community can justly take pride.

Jan 1, 1968

By Robert R. Dorfler

We ought not look back unless it is to de- rive from past errors, and for the purpose of profiting by dear bought experience. - George Washington During the mid 1960's, the application of automation was introduced to Climax by two process control companies. In 1969 Climax decided to seriously consider improved control of selected areas of their milling process after preliminary research had been completed on two rougher ball mill grinding circuits. In 1971 the Climax Mill Department developed a direct digital control scheme for rougher milling, designed a remote control center for handling process water, planned to automate the central steam heating facility for energy conservation, and convert other plant areas to automation. All twenty rougher grinding mills and the industrial water system were automated by 1977. The idea of automating central heating and other areas was abandoned although some fuel conservation items were installed at the heating plant. During this period from 1971 to 1978, the mill capacity was expanded from 39,000 metric tons per day to 43,000 metric tons per day and cyclone classification was introduced to rougher milling. Ore quality changes degraded the milling recovery from over 90% to less than 85%. This report describes the procedure the mill used in developing the automatic process and the occurances that affected the timing and fina1 control logic. The report deals primarily with the automation of the ball mill grinding, classification and flotation process as this was the bulk of the seven-year project.

Jan 1, 1979

By Alwyn Protheroe, Steve Smith, Herbert Jr. Gibson

In September of 2000, Cowin and Company, Inc., contracted with Paramont Coal Corporation to construct a large decline approximately 820 m (2,700 feet) in overall length at a grade of 8 degrees to the Lower Banner Coal Seam. This paper will discuss the methods used by the contractor and owner to overcome the unique problems required to construct this tunnel in a coal mine environment. The problems include, initial design considerations, equipment selection, ventilation, ground control, control of water in?ows, and placement of concrete.

Jan 1, 2003

By Carl E. Williams

Several years ago it was decided that ASARCO was in need of a good mine scheduling program to complete the computerized open pit mine planning package that was being used on new prospects and operating properties. A trial and error approach was being used to schedule a mine on a year by year basis, but this was time-consuming and far from offering an optimum solution l. Due to the magnitude of time and expertise required to develop such a sophisticated computer program, it was decided to search outside of the company for one already developed and available for purchase, This proved fruitless and therefore proposals were requested from consulting firms that had the expertise to develop such a program. Out of the five firms that submitted proposals, the one chosen was Systems Control, Inc., (S.C.I.) of Palo Alto, California. S.C.I. was selected because it was felt that they offered the greatest capability in the mathematical solution of complex problems. But due to their lack of mining expertise, a cooperative venture was developed whereby ASARCO would supply this mining knowledge. The selection of this vendor, the formulation of the contract and specifications, and the development of the computer program required about 3 years to accomplish.

Jan 1, 1973

By Frank J. Markewicz

The expanding: demand for ilmenite for use in the manufacture of titanium pigments, the knowledge that much of this raw material comes from distant points and therefore has to include a heavy freight charge in the price of the delivered product, and the further knowledge that ilmenite occurs both in the crystalline rocks of northern New Jersey and also in some of the Triassic igneous rocks, were considerations which led to a search for ilmenite-bearing sands in southern New Jersey. That this search was successful is now history, and it is the senior author's privilege to state here that most of the credit belongs to his associates who did all of the heavy work of hand-augering for samples, and who made all of the heavy-mineral separations. We all are grateful to members of the Department of Geology at Rutgers and Princeton Universities for the use of equipment which enabled us to evaluate certain samples from our ilmenite exploration program. In addition, we wish to express our appreciation to many companies for their cooperation and for their permission to publish mineral and chemical analyses. We also acknowledge with much appreciation our debt to the geologists and engineers with whom we have had many lively and rewarding discussions relative to the ilmenite deposits.

Jan 1, 1958

By B. S. Crocker

Introduction Pebble mill grinding can be defined as wet or dry grinding in a tumbling mill with pebbles used as grinding media instead of metal balls. To clarify the terminology, which may differ in the various mineral industries, it is essential to point out the differences between grinding with partially sized and carefully sized grinding media. When grinding with partially sized media (limited top size only), sizing is usually the setting of the primary crusher and there is no bottom sizing-all fines, chips, sand, and small sizes are included. This is often called run-of-mine or primary autogenous grinding. In this case the entire crusher discharge is fed into the mill and part of it becomes "charge" and the rest is "sand feed." When grinding with carefully sized media (i.e., media that is sized both at the top and bottom of the size range), no fines or chips should be present in the charge to the mill. This is pebble milling and the grinding media consists of ore, rock, Danish pebbles, gravel pebbles, or ceramic pebbles-media that is appreciably lighter than steel balls. The grind¬ing charge is fed to the mill to maintain a load and it does the grinding usually on a classified sand product of suitable size. If the grinding media is the same as the raw material being ground, this type of grinding is also called secondary autogenous grinding. Pebble milling was in common use in the gold mines of South Africa in the early 1900s. It is by far the oldest form of autogenous grinding. It may be used to grind crushed ores from a maximum size of about -% in. to all -10 µm. Pebble milling differs from "run of mine" grinding in many ways, but two big differences are: I) The ore itself is crushed to approximately t/4 in., or it may be further reduced to 8 or 10 mesh in a rod mill or fine crusher. 2) Then carefully screened ore of the correct size is placed in one or more stages of pebble mills to grind the ore to its desired final size. The media size is carefully selected to be the most efficient for the particular size reduction required in each stage. This is a precise practice, and pebble mills can be designed as easily and accu¬rately as can rod mills or ball mills by those skilled in the art. A pebble mill is a rotating cylinder in which the grinding media has been carefully sized for the job in hand. If the mill is doing coarse grinding, the media will be correspondingly large and the mill might be called a primary (or intermediate) pebble mill or a lump mill. If it is to do fine grinding, the media selected is proportionately smaller and the mill would be called a secondary or tertiary pebble mill. The media itself is normally rounded pebbles made from screened ore from the primary-crusher discharge. The charge could be rounded pebbles of the correct size removed from a primary autogenous mill, or the pebbles could be carefully sized gravel, sized beach pebbles, or sized flint pebbles. Usually the specific gravity of a pebble is from 35 to 55% of that of steel balls. It commonly varies from 2.66 to 4.2, depending on the specific gravity of the ore. Primary pebble mills, or lump mills, would be charged with pieces of rock up to 10 in. maximum and 5 in. minimum and would grind crusher discharge with a maximum size of 1/4 in. on hard ores and 3/8 in. on softer ores. Secondary pebble mills could be charged with pieces of rock screened on 4 x 4-in. square mesh and retained on 1 1/2 x 1 1/2 in. These mills would grind material with a top size of 8 to 10 mesh. Tertiary pebble mills could be charged with screened rock that would pass 3 x 3 in. and be retained on 1 x 1 in. and grind 35 to 48 mesh sands. Principles of Operation Media Size To operate a pebble mill efficiently it is imperative that great care be taken in screening the rock feed that is fed into the mill to round up into a range of pebble sizes that are suitable for grinding of the sands that are also being fed into the mill, having regard to the top size and the average size of such sands. Since most mill operators know from experience the correct size of steel balls to be used in a mill to handle different sizes of sand feed, it is convenient to relate the pebble size to a ball size. If the correct ball size is not known, it is relatively simple to conduct batch laboratory tests in 12- to 18-in. laboratory mills to determine the optimum size of steel ball. Also, Bond and others have published formulas for calculating the optimum ball size. When the correct size of the individual steel ball is known, then a pebble of the same weight will perform the same grinding if used in the proper size of mill and under the proper conditions, which will be defined later. This media size, as defined by the average weight of the individual pieces, is most important. The following tests will serve to clarify this point. Laboratory Tests Using Steel Balls or Pebbles. Several labora¬tory grinds were made to compare a given size and weight of steel ball with a pebble of the same weight but larger size, and of the same size but lighter weight. The following loads were used in the 12-in. laboratory mill: All 1 1/4-in. steel balls weighing 122 ± 10 g. All 1 3/4-in. pebbles weighing 123 ± 5 g. All 1 1/4-in. pebbles weighing 45 ± 3 g. To equalize capacity, the test with the steel balls was run only 34% as long as the tests with pebbles. In each case, 2,500 g of primary sands were ground at the optimum gravity for the media being tested at a speed 78 1/2 % of critical speed, and for 170, 340, and 680 revolu¬tions for the steel balls, and 500, 1,000, and 2,000 revolutions for the pebbles. The first test is summarized in Table 74 by showing the 80% passing size in microns for the feed and for each of the grinds. Two other tests are summarized in Table 75 showing the screen analyses

Jan 1, 1985