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By David G. Chedgy

Noise control is an important consideration in the design and operation of coal preparation plants. It is essential to understand and control noise sources to provide a safe environment for operating personnel. The average noise exposure levels in a typical coal preparation plant generally range between 95 and 100 db. An operator should not be exposed to noise levels of this magnitude for more than 4 and 2 hr, respectively, according to MSHA (Mine Safety and Health Administration) regulations. The major noise sources in a coal preparation plant include noise from impacting the inside surfaces of steel chutes, vibrating screens, coal crushers, pumps, motors, fans, and blowers. The magnitude of noise varies considerably from one situation to another. NOISE EXPOSURE IN COAL PREPARATION PLANTS Noise regulations established by the Occupational Safety and Health administration (OSHA) are accepted by MSHA relative to caal preparation plants. Fig. 14-43 shows a graph of the time of noise exposure in hours per day to which a worker can be exposed for each A-weighted sound level. According to this graph, a worker is permitted to spend 8 hr in a 90-db noise environment, but that time must be halved when the A-weighted sound level increases 5 db. The noise dosage a worker receives is determined by the ratio of the length of time the worker spends in a particular noise environment divided by the length of time he is permitted to spend in that environment. If the worker is exposed to several different sound levels, his total dosage would be the sum of each of the individual dosages. The equation for determining the dosage is: where C is the actual duration of exposure at a given steady state noise level and T is the noise exposure limit for the level present during the time C According to MSHA regulations, the dosage, D, should not exceed one (1) for any worker for a full day of work. As an example of typical plant noise levels, vibratory screens and crushers emanate noise levels in excess of 90 db at a distance of 3 ft (0.9 m) with no material flow. With material flow, noise level increases 1 to 10 db depending

Jan 1, 1991

By Ben F. Dickerson

For a number of years in office bull sessions, I have not so jokingly discussed the engineering or management fiasco of the year or decade. Usually, governmental projects are in the vanguard, such as tanks that won't tank, or heavy water plants that don't produce heavy water, etc., etc., all to the flutter of the taxpayers' hundreds of millions, or billions of dollars. The mining industry has also brought forth a share of this type of fiasco. Ordinarily, due to lack of funds, not opportunity, the stockholders only stand losses of a few tens of millions. Little knowledge of these major errors gets to the outside world, for the ordinary stockholder is usually not very alert, and managements are extraordinarily adept at discreet midnight internments in the fine print of the footnotes of annual reports.

Jan 1, 1975

By Guy T. McBride

Mr. Chairman, Fellow Keynoters, Ladies, and Gentlemen: It is purposed to consider the theme of this meeting, "The 1970's - Pivotal Decade for the Mineral Industries," from the particular -- some would say peculiar -- point of view of one engaged in mineral engineering education, and this coign of vantage prompts the textbookish subtitle, "Problems and Solutions (?) in Undergraduate Mineral Engineering Education." Three interrelated problems, or problem areas, will be presented and discussed, and where solutions seem to be emerging or seem worthy to be suggested, they will be identified. The three problems are: 1. The goals, and therefore the course requirements, of the undergraduate curricula are in question, 2. The unit costs of undergraduate education appear excessive to those who bear the charges, and 3. The social ambience of the day affects for the worse the work attitudes and the career motivation of both prospective and in-course students. The Need. - Any discussion of these problems must bottom upon recognized national needs for minerals and a technologically sound minerals industry staffed by competent engineers, particularly at the exploration, development, and production levels, Fortunately, these needs have been recognized and to some extent quantified, and they are even now under further intensive study by some of the most competent members of our society and profession. The latter reference is, of course, to the current work of the National Commission on Materials Policy (U.S.A.) under the executive directorship of Dr. James Boyd. Predating this, and perhaps to some extent

Jan 1, 1972

By G. E. Erdman

The total attitude of the industrial minerals producing company is a most important factor in how they regard quality control, technical support for their customers, and procedures used in mining their minerals products. This philosophy of the producing organization directs their action along certain courses. For instance, one producer might see itself as a mining company with its major goal mining efficiency to make it exceedingly price competitive in the market place. Another producer might see itself as a wide spectrum multi-mineral producer that furnishes a reputation for a complete line of products. Another producer might see itself as a specialist in one specific industrial mineral and bet its whole bankroll on providing the very best of quality with the goal of becoming the largest producer of the mineral commodity. Each perception can have a different impact upon the producer's operations and its regard for quality control. Minerals Being Produced The specific markets being served will have their own influence on how mineral product quality is regarded. Ceramic raw materials, for instance, have a very low tolerance for iron which is the cause of low fired color. Thus, industrial minerals for use in glass and ceramic markets must be won from deposits with a very low iron content. A high iron silica sand might be sold to a manufacturer of good old beer bottles. That is a solution to one particular quality facet, but eliminates selling to a significant market in clear glass, porcelain enamel, ceramics whiteware bodies and hobby glazes. Thus, there is a need for sound research in the industries expected to be served before the product is brought to market.

Jan 1, 1987

By S. Nowosad, O. Langefeld, A. Godge

Pumped storage power plants are one of the options to store the fluctuating production of electric energy from renewable sources in a sustainable manner and is able to provide Europe with the opportunity to overcome the effects of climate change. However, due to the extensive land use required when planning the water reservoirs, several projects have been declined after social conflicts raised, even in Germany. Nonetheless, after the flooding and drought periods in northern Germany, the catastrophic floods of July 2021 in south west Germany, the country’s need to decrease its dependency on import of gas and energy supply, and the energy transition goals set for 2030 and 2050, it has become clear that the implementation for water reservoirs must increase in the coming years, whether above ground or underground. Furthermore, not only Germany is interested in the use of underground cavities for water and energy storage, other European countries as Sweden are currently developing studies and solutions to bring this concept soon to reality in Europe. This article introduces the systematic approach for planning and designing large underground cavities for water reservoirs as a sustainable energy transition approach and a climate change response by showcasing the process of selection and results of the “Energy and Water Storage Harz” (EWAZ) research project. Additionally, the results of the evaluation of the effect of implementing the selected underground and surface infrastructure into the current water management system in the region is presented. Finally, the different possibilities evaluated for the underground cavities within the project and the technical challenges involved are be discussed.

Feb 1, 2023

By Bradley Johnson

Introduction Since its creation in 1910, the Bureau of Mines has conducted a wide-ranging program of research on mining technology. Over the years, the Bureau has perhaps become best known for its accomplishments in mine health and safety technology, a field that has been the object of much public attention. Nevertheless, the Bureau's program in other areas of mining technology has always been strong and remains so today. A distinguishing feature of the current effort is its tendency to concentrate on high-risk, high-gain projects. Throughout the federal research establishment, the Administration's policy has been to phase out short-term research that could be performed by private industry, in favor of long-terms projects that industry cannot usually afford to fund. This tendency is evident in the following account of the highlights of the Bureau's current mining technology research program. Underground Mining Technology Fragmentation. Research studies on fragmentation have focused on nonexplosive concepts. If successfully developed, these concepts would lend themselves to continuous automated operation and reduce time and costs of fragmentation in underground mines. A special laboratory-scale device to investigate ripper cutting was designed and fabricated in 1982. Experiments were conducted with large drag cutters to develop an efficient cutting technique for hard rock. Tests were successfully conducted with 76- and 152-mm (3¬and 6-in.) bits in rocks up to 138-MPa (20,000-psi) compressive strength. Because of the success of these initial tests, larger-scale proof-of-concept tests are planned. To prepare for these tests, a tester capable of cutting a 1.8-m (6-ft) opening was fabricated. This device will allow ripper cutting to be carried out at full scale, and will also allow mucking and holddown requirements to be investigated. Preliminary estimates indicate that this concept has the potential to double excavation rates while reducing costs up to 30%. Mobile Hydraulic Monitor. A mobile hydraulic monitor system was designed, fabricated, and used to mine the St. Peter sandstone at an average rate of 22 t/h (25 stph) from a face in Martin-Marietta's foundry sand mine at Clayton, IA. A slurry transport system worked in tandem with jet cutting, resulting in a continuous mining and transport system. Problems with spray from the jet hitting the sandstone showed that a remote-controlled mobile monitor was needed at a station 15 m (50 ft) from the face, so the operator would have a clear view of the face and be safely out of the backwash. Crushers. Research on crushing systems to improve efficiency in underground mines is continuing. Final testing of a low-profile, portable crusher for underground applications has begun. A system for handling oversize feed to the crusher was conceived and tested. Underground crushing systems, portable enough to be kept near the face, offer considerable improvement in materials handling because main haulage by conveyor is often less costly. Hoists. Research to increase hoisting efficiency and speed and to reduce energy requirements has led to experimental ore skips. The skips feature guide rollers to improve riding characteristics, bearing improvement at critical locations, and weight savings of about 1.8 t (2 st). An experimental aluminum skip was installed last year at Kerr McGee's Church Rock mine. Performance is being monitored under production conditions. A portable service hoist for low-cost, efficient movement of men and supplies in raises between mine levels was designed, fabricated, and tested on the surface. The hoist has maximum equipment payload of 570 kg (1,250 lb) at

Jan 11, 1983

By S. T. Hall

The authors have previously described a method of enhancing the magnetic susceptibility of pyrite by a mild alkaline oxidative pressure leach (Hall et al, 1985). This finding was successfully applied to enhance the magnetic desulfurization of an Eastern Canadian coal (Hall and Finch, 1984). This paper reports on enhancement of the magnetic susceptibilities of iron-containing sulfides (arsenopyrite, chalcopyrite and marcasite) upon similar treatment. It also explores the application of this pretreatment in the magnetic separation of fine sulfides from oxides and other sulfide minerals. In the latter case galvanic interference effects have a serious impact on the potential use of this method.

Jan 1, 1985

By Suresh K. Bhatt

A mathematically simulated modeling technique is used in this U.S. Bureau of Mines paper to represent a hypothetical mining operation with existing mining technology and prevalent mining costs. Mining scenarios are prepared and evaluated for potential benefits and costs available through computer-assisted mining. Base criteria, parameters, and methodology are described.

Jan 1, 1991

By G. S. Dobby

The Cu-Au flotation cleaning circuit at Freeport Indonesia consists of three stages: two flotation column stages, in series, followed by one stage of mechanical cells. Concentrate from both column stages is combined as final concentrate. The mechanical cell stage has two parallel rows of seven Wemco 500's. The Wemco concentrate is recycled to feed of the second column stage and Wemco tails report to final tailings. This paper examines cleaner circuit performance, focusing on size-by-size performance of the individual stages and the individual mineral groups -copper sulfides, gold and in sol minerals. The recovery mechanisms of insol minerals are examined.

Jan 1, 1995

By N. O. Lotter

Studies investigating the evaluation errors in laboratory scale flotation tests for platinum-bearing Merensky Reef in the Bushveld Complex, South Africa, have constructed a quality control model to assure a 95% confidence in the attested data. Specialized knowledge of the mineralogy of the platinum-group elements (PGE), together with acquaintance of evaluation models, validate a systematic checking of the sub-processes of the testwork. These include determination of minimum mill feed sample size, assay head reference distribution, mass balance checks, and specialized sample preparation of flotation test products prior to analysis. A flow schematic of this quality control model is proposed, with an example of test data that produce an error in the metal balance of 0,2% after rejection of outlier data.

Jan 1, 1995

By A. B. Cecala

The US Bureau of Mines has designed and evaluated total mill ventilation systems (TMVS) at two mineral-processing operations. Both systems have proven proficient at cost-effectively reducing respirable dust levels throughout the mills. An 11.8-m3/sec (25,000-cfm) system installed at a clay-processing mill provided about 10 air changes per hour(acph). This system reduced respirable dust concentrations by about 40% throughout the mill building. The second evaluation was performed at a silica sand operation. Tests were performed with 23.6 and 47.2 m3/sec(50,000 and 100,000 cfm) of ventilation to the mill building, corresponding to 17 and 34 acph. Average mill-wide respirable dust reductions were 36% and 64%, respectively. This work was done to determine general guidelines and design criteria for operations that may be interested in implementing TMVS. TMVS should be designed to draw clean make-up air in near the base of the mill. This air provides general purging and may clear some dust-laden areas throughout the mill structure. It is discharged at or near the top of the structure. Without the TMVS, dust can exit the building through open doors, windows, etc. and contaminate outside plant workers. This dust can also be recirculated back into the mill building causing additional contamination problems. TMVS is not a stand-alone system. It is a supplemental technique to assist other localized dust control systems (Mody et al.,1987).

Jan 1, 1995

By R. E. Johnson, J. E. Dutrizac, B. R. Palmer

Silver sulphide (acanthite) and lead sulphide (galena) dissolve relatively rapidly in FeCl3 media with the formation of elemental sulphur and soluble silver and lead chlorides. Parabolic kinetics are observed, and the leaching rates are controlled by the outward diffusion of the AgCl and PbCl2 reaction products through the constantly thickening elemental sulphur layer formed on the sulphide. Hence, both leaching rates are strongly dependent on those variables which affect the solubility of AgCl or PbCl2 at the reaction interface. The rates therefore increase with increasing temperature and total chloride ion concentration, but decrease in the presence of elevated concentrations of AgCl or PbCl2. By contrast, neither sulphide is significantly attacked by ferric sulphate media, even at high Eh values. Although the reasons for the refractory behaviour of Ag2S in sulphate media remain unknown, the low rate of PbS attack is attributable to the passivating layer of PbSO4 and elemental sulphur formed on the surface of the galena.

Jan 1, 1993

By M. Tomono

Harima Works is operating as the zinc and lead simultaneous smelter utilizing the imperial Smelting Process. The smelter commenced its operation in 1966 with designed capacities of 3500 t/M of zinc and 1500 t/M of lead. After that, owing to the results of progressive developments and improvements in operating efficiency and equipments, it has doubled its production figure. This report reviews our recent developments carried out from 1983 to 1985. Firstly, the furnace bottom design was modified for the purpose of reducing carbon consumption ratio. As a result, shaft operating condition has been satisfactory and carbon consumption ratio was improved by 2%, in addition, furnace zinc production was increased by about 8%. Secondly, the oxygen enriched operation at sintering machine was tested from July 1984 to June 1985, in order to increase the sinter productivity, and to solve the capacity imbalance between sinter plant and ISF. As the test result, it was found out that sulfur elimination rate was increased by approx 10% with the oxygen enrichment of 500 Nm3/hr.

Jan 1, 1986

By G. L. Mowrey

One of the critical parameters of any automated coal mining system is the ability to only cut coal and not the rock / waste surrounding it. To do this the system must be able to detect where the coal-rock/ waste interface is. The U.S. Bureau of Mines is investigating a new approach to vibrational coal interface detection (CID) by utilizing an adaptive signal discrimination (ASD) system to perform feature extraction and classification of complex vibrational signals. Accelerometers, geophones, and piezoelectric films are mounted on mining machines and/or strata and are used to monitor vibrations produced during mining operations. Selected data are digitized and the ASD system is trained on known types of vibrations (e.g., "cutting coal" and "starting to cut roof"). A signal classifier is then generated which can be used to identify unknown waveforms. The ASD system can be retrained if mining conditions change significantly.

Jan 1, 1990

By J. Hahn

Modern high-performance disc filters such as the BoVac Disc Filter have set new standards in applications where large capacities of solids, liquids and air have to be managed through the filters. These disc filters have an enormous hydraulic capacity. For applications in mineral processing - with a special focus on tailings – which typically have high solids throughputs but low filtrate and air flows, this capacity can be used to expand the filtration area while maintaining the benefits of very good cake discharge even with thin cakes (4-5 mm) and a high filter speed. Research work over the last century has proven that increasing the filter area should be achieved by increasing the disc diameter. Because adding more discs leads to big problems with the cake discharge. However, as the disc diameter increases, the filter segments become longer and heavier. A suitable compromise was therefore found by doing both. The result is the first vacuum disc filter that offers significantly more than 300 m² of filtration area, but has no more than 6 discs.

Feb 1, 2023

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

9.1 INTRODUCTION There are approximately 70,000 abandoned coal mines nation¬wide, which is about 35 times the number of underground coal mines presently operating. The US Bureau of Mines estimates that there are over 8 million acres of undermined land due to the extraction of coal, metals, and nonmetals. Subsidence has affected more than 2 million acres, and more than 99% of the subsidence is related to underground coal mining. There is reason to believe that some of the remaining 6 million acres of the undermined land have a high potential to subside. The expansion of housing, highways, commercial structures, and other facilities has required the use of many areas that are underlain by abandoned coal mines, and this growth will continue. Many subsidence problems are derived from the collapse of aban¬doned underground coal mines. There are numerous abandoned mine workings in the anthracite fields of northeastern Pennsylva¬nia, in the Appalachian bituminous fields, the Illinois Basin, the Rock Springs, Wyoming area, and other areas of the United States (Gray et al., 1976). Various room and pillar patterns of mining have been used in the dipping anthracite seams and the nearly flat-lying bituminous seams with considerable variation in the per¬centage of coal extracted. The progressive deterioration of pillars, mine floors, and mine roofs after long exposure to air and water may later result in the collapse of strata over the mine entries, the crushing of the remaining coal pillars, or the bearing failure of the mine floor beneath the coal pillars. Subsidence then results as the collapse reaches the ground surface in the form of differential strains, depressions, cracking of the ground, and sinkhole devel¬opment. Subsidence over active longwall mines, which occurs concurrently with mining or is completed within a short period following coal extraction, has been studied extensively over the past decade. On the other hand, subsidence over abandoned coal mines receives little attention by the researchers, mainly because it is difficult to predict and takes place decades after mining has ceased. The techniques of investigating the subsidence events over abandoned coal mines are similar to those employed for active mines except that at the outset, it is necessary to determine whether or not the subsidence events are mining-related (Chugh et al., 1986; Cummings and Singh, 1986; Peng and Hsiung, 1986). This calls for the identification and confirmation of abandoned mine workings under or near the affected surface structures. Gen¬erally old mine maps, if available, are acquired, the surface bore¬holes are drilled for confirmation of the accuracy of the mine maps and determination of the potential for continued subsidence. Sub¬sidence monuments are established and periodic surveys con¬ducted to determine the amount and trends of surface movement in and around the affected surface structures; surface boreholes are used to investigate the integrity of the underground structures (i.e., roof, coal pillars, and floor) by TV camera. They are also used for monitoring the vertical and horizontal movements of the subsur¬face strata by Sondex (FPBX) and inclinometer (PFBI), respec¬tively, for determining the continuity of the subsidence events. Tape extensometers, crackmeters, etc., are used to monitor the development trends of major cracks in the structures or on the ground. These data are used to identify the causes of the subsidence events. Finally, abatement methods are selected to stabilize the structures. It must be noted that most subsidence events over abandoned coal mines are reported after the fact, and investiga¬tions are begun some time after reporting, that the subsequent measured movements are generally much smaller than that at¬tained in the active mines, and that due to lack of knowledge about the damage conditions and the exact location of the abandoned mine workings with respect to the affected surface structures, the precise causes of surface subsidence or surface structural damages are in most cases very difficult to identify. 9.2 TYPES OF SURFACE SUBSIDENCE According to Gray et al. (1977), after examining 354 incidents of subsidence above abandoned mines in the Pittsburgh metropol¬itan area, the subsidence features have a mean diameter (i.e. the average of long and short dimensions) from less than 1 ft to 1600 ft, with 84% less than or equal to 15 ft; the subsidence features have a depth ranging from less than 1 ft to 48 ft, with 89% less than 25 ft; 66% of the subsidence features are deeper than they are broad. Nearly 59% of the subsidence features occur with overbur¬den less than 50 ft thick and 81% less than 100 ft thick. No subsidence features occur with overburden thicker than 450 ft (Fig. 9.1). Occurrence of subsidence incidents varies from imme¬diately to more than 100 years after mining (Fig. 9.2). In analyzing the characteristics of approximately 3000 chim¬ney subsidence features along the Colorado Front Range, Matheson and Eckert-Clift (1986) examined historical aerial photographs on 4- to 14-year intervals between 1937 and 1967 and found that the majority of observable surface subsidence features occurred within 30 to 40 years after mining. According to Gray et al. (1977), the most prevalent subsi¬dence features over abandoned mined land are sinkholes, with depths of more than 3 ft, and troughs or sags, usually less than 3 ft deep. Sinkholes are steep-sided pits, while troughs are shallow depressions much wider in area than sinkholes. A. SINKHOLE SUBSIDENCE A sinkhole is caused by the collapse of a mine roof that works its way upward. If it is not arrested during the process it will eventually reach the surface and emerge as a sinkhole. The process is governed by the thickness and character of the overburden, the width and height of the mine openings. Sinkholes are usually 3 to 20 ft deep and may be 2 to 40 ft in diameter, although most are fewer than 16 ft across (Gray et al., 1977; DuMontelle and Bauer, 1983). Newly formed sinkholes have steep sides with straight or bell-shaped walls. At times, they appear to be conical in profile with the apex upward. If the topsoil collapses, the top portion will widen to form an hour-glass shape. Sinkhole subsidence usually occurs over abandoned mines less than 165 ft deep (Hunt, 1979). Matheson and Eckert-Clift (1986) found that chimney sink¬holes are likely to occur when the ratio of overburden thickness to mining height (h/m) is less than 4 to 5. When h/m is between 5 and 10 to 11, the potential occurrence of chimney sinkholes decreases rapidly. When h/m is more than 10 to 11, less than 10% of the mine openings that collapse will induce sinkholes on the surface.

Jan 1, 1992

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

By S. Usui

In a previous article1 the methyl stretching vibrations for a series of xanthates and amines were reported, and the frequency of the methyl stretching vibrations in a homologous series of the alkyl compounds were shown to decrease exponentially with chain length and approach a limiting value corresponding to that of saturated hydrocarbon beyond six or seven carbons. These observations were explained quantitatively by the theory of Goto and Takenaka2 who proposed an equation relating the frequency shifts to the inductive effect of polar groups. As a result, the alkyl groups in the shorter chain heteropolar compounds were interpreted as becoming somewhat polar, and a good correlation was observed between the CH3-stretching frequencies and such surface chemistry parameters as contact angle, surface tension lowering, solubility, and critical micelle concentration" The substitution of fluorine atoms, the most electronegative of all atoms, for the hydrogen atoms in alkyl groups was expected to reverse the direction of the inductive effect within a heteropolar molecule, and its effect on the C-F stretching frequency shifts and on some surface chemical properties became of interest. In a recent study the C-F stretching frequencies of a series of normal perfluorocarboxylic acids and a few 1,1' -dihydroperfluoro alcohols were determined as a function of chain length and the results were compared with the corresponding hydrocarbon analogs. Furthermore, the nature of the induced polarity in the hydrocarbon and fluorocarbon groups was investigated through surface potential measurements at the interfaces of air-aqueous solutions of normal butyl alcohol and of l,l'-dihydroperfluorobutyl alcohol and through differential capacity and electrocapillary measurements at the interfaces of mercury-aqueous solutions containing the above two alcohols. Experimental Normal perfluorocarboxylic acids and 1,1' -dihydroperfluoro alcohols were provided through the courtesy of the 3M Company, St. Paul, Minnesota. Normal butyl alcohol was purchased from Distillation Product Industries, Rochester, New York. A Perkin-Elmer Model 621 infrared spectrophotometer was used to investigate the spectra in the frequency region of 1200 to 1270 cm-l with frequency scale

Jan 1, 1971

By G. A. Kordosky

Solution mining followed by solvent extraction-electrowinning (SX-EW) is one possible way to economically extract copper from ore bodies which cannot otherwise be profitably exploited. Published costs for the unit processes of SX and EW are out of date especially in view of the newer reagents now available and the newer circuit configurations, This paper updates the costs for SX-EW in 1985 dollars and also discusses some of the trade-offs which must be made when designing an SX-W plant.

Jan 1, 1986

By C. A. Rowland

With the rapidly rising cost for electrical energy and the long range pre¬dictions for continued rising costs and even possible energy shortages, operation of rod, ball, autogenous and partial autogenous grinding circuits to efficiently utilize the power delivered to the mills will be demanded. This calls for a method to evaluate grinding circuit performance that is accurate, reliable and readily usable as an operating tool by supervisory, technical and operating personnel. Neither the Rittinger and Kick theories of comminution,(1) which preceded the Bond Theory by more than 50 years, had a mathematical means that could be used to predict and evaluate the performance of crushers and grinding mills used to comminute ores and rocks. This severly limited the practical use of these two theories, which were contradictory to each other. They have been superseded by the Bond Third Theory of Commnution.(1)

Jan 1, 1976