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By Adrian C. Dorenfeld

Laboratory data form the basis for the design of flotation circuits. These data, obtained from testing samples of the ore, should show the optimum con- ditions for concentrating the ore and the effects of change in process vari- ables. Among the more important variables commonly met in flotation practice are: Liberation Size-This is the particle size at which substantially all of the valuable minerals are detached from the gangue minerals. Reagent Additions- Experimental data will show the reagents used and their quantity, quality, and point of addition. Pulp Density-This is an important variable in the design of the flotation circuit because it is one of the factors determining the size and number of flotation cells. Flotation Time-The time necessary to make the separation into concentrate and tailing depends on such factors as particle size and reagents used and must be known for determination of the size and number of flotation cells in the plant. Temperature of Pulp-Temperature affects flotation reactions. Water at approximately room temperature is satisfactory for most flotation separations, but some minerals such as fluorspar are best floated in hot water. Type of Circuit-Many types of circuits can be designed, as discussed in detail later. The laboratory tests should provide data for design of the plant circuit best suited to the ore. Water-Knowledge of the quality of the future plant's raw water is important because water quality may drastically affect flotation results. If possible, flotation tests should be made using samples of the water to be used in the plant. Flotation results may vary with seasonal temperatures and with circulation of reclaimed water. Experimentation can show whether it will be necessary to provide for heating, cooling, treating, or reclaiming the water. Uniformity of Ore-The extent of variation of the ore in hardness, grind- ability, mineral content, and floatability should be investigated so that the designer can ascertain methods to overcome problems caused by ore variations. Settling and Filtration Data-To design the plant flotation circuits, settling and filtration rates and ultimate percentage solids of the various products must be determined. Corrosion and Erosion-Corrosion and erosion tests should be made on the flotation pulps, particularly if the pulps are acid or at an elevated temperature, so that suitable materials of construction can then be chosen.

Jan 1, 1962

By F. Lynwood Garrison

The study of the microscopic structure of the iron of car-wheels, which it is the aim of this paper to describe, was made at the suggestion of Dr. Dudley, whose paper upon the constitution of cast-irons (page 795 of the present volume) should be read in connection with this one. Although the two wheels referred to in that paper possess nearly an identical ultimate chemical composition, they possess, as Messrs. Dudley and Pease have shown, physical properties so widely different, as to make one of them nearly worthless for the purpose for which it was cast. Considering the chemical similarity, such a marked difference in strength cannot but be very striking, and goes to show how little reliance can be placed upon chemical analysis alone, if not properly interpreted. I say " not properly interpreted," because I believe that very few analyses of iron or steel are thus fairly dealt' with. That an analysis may read so much of one thing and so much of another, tells us very little as to the properties of the metal; for these are determined to a large extent by the conditions and relations in which the elements are present. To the unaided eye the difference between the two wheels under consideration is not striking. The inferior one would be described as light gray, and the other as somewhat darker. Upon careful examination of an etched fragment, however, the difference in structure will be found to be very marked, as may be seen in Figs. 1 and 2 (produced from photographs by the Ives process). It is exceedingly difficult to produce a photograph giving a fair notion of the appearance of the metal under the microscope, so as to give to persons not familiar with microscopical study, a proper appreciation of what a microscopist would consider a decided difference in structure. I would, therefore, urge any one especially interested in this subject, not to depend upon the photographs of others, but to prepare and etch the specimens himself, and then examine them with a good microscope. As is well known, iron possesses the property of uniting with a number of elements, forming products which are either highly intimate mixtures of several substances, or else compounds of an indefinite character. Indeed such mixtures may be regarded as alloys, or as Matthiessen puts it, " solidified solutions of one sub

Jan 1, 1886

By R. Dawes, B. R. Davis, T. LeBel, C. Pelow

Settling of copper, lead, and other metals in slags is an important part of overall recovery of metals, both in primary and secondary operations. The impact of slag reduction on settling is not well known. This is important since the cycle time for slag cleaning can be significantly changed with reduction of slag, through a decrease in slag viscosity and growth of metal/sulphide particles. KPM has worked with a number of primary and secondary recyclers to better understand how reduction of slag can improve overall operation. In this presentation, we describe the type of test work and analysis done and how this information can be used to guide improvements in metal recovery from slags. INTRODUCTION Typically, copper smelter slags contain about 1-2% copper, both in the form of copper and copper sulphide beads and dissolved copper oxide in the slag (converter slags are recycled to the smelter). With the average grade of world copper deposits close to 0.7%, it is practical to recovery copper from these slags. This is done routinely by one of two methods. The first method uses slow cooling in large ladles or beds to grow particles of copper sulphide. The slag is then crushed, milled, and floated for copper recovery. The slow cooling promotes precipitation and growing of the sulphide particles from the solution and then, the slag can then be milled to recover the copper through froth flotation. A consideration of this process includes the energy cost related to grinding, the use of water, capital cost for ladles, and the availability of tailing dumps. At the same time, the retreatment of the Cu concentrate produced in the mill will displace fresh concentrate, which ultimately results in a reduction of the throughput. The second method is the so-called “slag cleaning” and is based on the settling of copper and copper sulphides particles dispersed in the slag. This is achieved by decreasing its viscosity by increasing the slag temperature and reducing the iron spinel solid phases (magnetite) that increase the apparent viscosity of the slag. Typically, the reduction does not extend to the reduction of copper, and soluble copper (mainly as oxide) and fine matte particles remain at concentrations typically at or above 0.5% (Cardona et al., 2011; Cardona, 2011). This is an improvement on 20 years ago when slag was sent to dumps at 1% or less (Demetrio et al., 2000).

Jan 1, 2019

By C. R. Fondersmith, John Chipman

In the manufacture of rimming steel—which constitutes the bulk of the world's production of mild steel—the factors that determine the quality of the product are very closely associated with the pouring and solidification of the steel in the ingot mold. The behavior of the metal in the mold is, on the one hand, the resultant of the complex processes through which the liquid metal has passed, and, on the other, it is an indication of the probable behavior of the ingot during subsequent rolling and finishing operations. A detailed study of the mechanism of solidification of rimming ingots ought to yield information that will be useful in controlling the structure of the ingot and the quality of the ultimate product. The first step in such a study is to determine the rate at which the solid shell of the ingot forms from the liquid metal in the mold. A knowledge of the rate of solidification will make it possible to correlate observations of rimming action and gas evolution with studies of the structure, porosity, and segregation within the ingot. Thus, if the metal in the mold is observed to fall, let us say one minute after pouring, the portion of the structure associated with this fall will be located in the part of the ingot that is solidifying at that particular moment. The knowledge that the ingot solidifies to a depth of 3/4 in. in the first minute enables us to locate in the ingot the structural features corresponding to the observed fall. Conversely, the blowhole condition associated with a rising or growing ingot will be found in the part of the ingot that is solidifying at the time the phenomenon occurs. The time required for the complete solidification of an ingot has long been a matter of speculation. It is of considerable practical importance in connection with charging and heating in the soaking pits. Too rapid handling of the ingot through heating and rolling processes has been thought to be associated with excessive pipe and other internal defects, so that a further object of this investigation was to attempt to estimate the time required for complete solidification, or the minimum practical time within which the subsequent operations can be performed with the minimum hazard.

Jan 1, 1937

By J. Ji, A. Shi, B. Grcic

Floating copper minerals selectively against iron minerals such as pyrite has been challenging for most copper and copper/gold deposits, and many operations continuously strive to maximise copper recovery whilst meeting target concentrate grades. Maximising grade is achieved by selectively floating copper minerals and rejecting pyrite. This challenge is to minimise copper and gold losses in the roughers if this approach is used. A flotation testing program was conducted by Bureau Veritas Minerals on a high-level pyrite bearing copper/gold ore sample to establish the potential to reject pyrite in the rougher stage, but also maximise copper and gold recoveries, and ultimately final product grades. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. The main variables tested for achieving selective flotation of copper and gold were collector type and pH control, through lime addition. Grind sizes used in the rougher and cleaner tests had been established from previous test work. The collectors tested were SIPX, Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The two best performing collectors were then tested in open-circuit cleaner tests, with the optimum conditions examined further in locked cycle tests. INTRODUCTION Bureau Veritas Minerals Metallurgical Division was retained by Eldorado Gold Corporation to perform a metallurgical testing program on a high-level pyrite bearing copper/gold ore sample from a gold- copper porphyry deposit to optimize flotation reagent scheme to improve gold and copper recoveries and copper content in the concentrate. The test program included feed characterization, rougher and open-circuit cleaner flotation as well as locked cycle flotation tests. Initially, various collectors were screened in kinetic rougher tests, and over a pH range of 8.5 (i.e the natural ore pH) up to 11.5, using a primary grind size of 80% passing 120 µm. The collectors tested were SIPX (sodium isopropyl xanthate), Aerophine® 3418A, Aerophine® 3421, Aerophine® 3422, Reagent S-11192 Promoter, Reagent S-11193 Promoter and FloMin C4150. The enrichment ratio, which is the ratio of copper to sulfur in the concentrate to that in the feed, was used as proxy for rating selectivity (i.e. degree of pyrite rejection or depression).

Jan 1, 2019

By J. Kalynchuk, R. Poulin, R. Pakalnis, D. Peterson

"A market survey of all underground mines in Canada (Fig. 1) was completed to determine the ex1ent to which cable bolts are being used as ground support. One hundred and nine questionnaires were distributed with 65% of them being returned completed. As a result, the quantity, types and locations of cable bolt installations have been determined. In addition, the grout pumps used, time required and cost of installations have been addressed. The popularity of other support types used in underground mines is summarized.IntroductionRock reinforcement consists of installing support within the rock mass (internal) or within the excavation (external). Within the past twenty-five years several new types of ground support have been introduced within the mining industry. Cable bolts were first introduced to the Canadian mining industry in the 19605 (Gramoli, 1975).Cable bolts have gained increased acceptance as a method of ground support during the past decade. The cable bolt is comprised of one central wire surrounded by six slightly smaller outside wires and typically having an over-all diameter of 15.2 nun to 15.9 nun (0.6 in. to 0.625 in.). The cable wires are made from high-strength steel with an ultimate strength of 26 tonnes (58 000 Ibs) (Goris, 1990). Current cable bolting practice consists of cement grouting one or more steel cables within a drill hole in order to reinforce the surrounding rock mass. Cable bolts are a form of internal support and are generally used in open stoping and cut-and-fill mining methods (Fig. 2).They can be either tensioned (active support) or untensioned (passive support) and may be of any length, though for underground mining applications, they are generally less than 20 m (65.5 ft) and are untensioned. A major advantage compared to previously used ground support is that the cables are flexible and can be coiled to approximately 1 m (3.3 ft) diameter coils allowing relatively long support to be installed in areas with limited head room (1.8 m).With the advent of bulk mining methods which require comparatively large stopes, it has necessitated the use of relatively long support tendons to stabilize the large spans being created. As a result, the use of cable bolts for ground support in mines has increased significantly within the past decade."

Jan 1, 1995

By J. T. Berling, J. B. Conway

MANY studies have been devoted to the correlation of low-cycle fatigue behavior. Several of these have also been concerned with the identification of an effective approach to the prediction of low-cycle fatigue behavior from short-term tensile data. Perhaps the first such approach was that reported by coffin1 in the form: where is the plastic strain range, Nf is the number of cycles to failure, and c is a constant characteristic of a material. In this analysis by Coffin it was reasoned that in a short-term tensile test the value of Nf could be considered as equal to a , in which case the value of c is interpreted as the tensile ductility, f, measured in such a test. If, therefore, the Coffin equation is assumed to be applicable to a given material, it is a simple matter to predict the low-cycle fatigue behavior. Once the tensile ductility is known, this point is located at Nf = 1/4 on a log-log plot of vs Nf and a line is drawn through this point with a slope of to yield a prediction for the fatigue life. Actually, while this equation is quite useful it is not generally applicable, for experiment has shown that the value of —$ does not reflect the relationship between and Nf for all materials or even for a given material over all temperature ranges. An equation similar to the Coffin equation was prepared almost simultaneously by Manson.2 This expression had the form: In this case the exponent on Nf was not assigned any particular numerical value in accord with the various values of this exponent which had been observed for different materials and at different test conditions. Because of this characteristic it will be recognized that little use can be made of this expression in predicting low-cycle fatigue results. A fairly effective expression for use in predicting fatigue behavior was proposed by Manson3 based on a study of numerous sets of experimental results. This relation (Method of Universal Slopes) makes use of modulus of elasticity, ultimate strength, and tensile ductility, and assumes an exponent on Nf equal to 0.6 in representing the plastic strain range contribution to total strain range. While this equation of Man-Son does not yield completely accurate results it must be recognized as being the best means available at

Jan 1, 1970

By Nicholas J. Grant, John Chipman

A FULL understanding of the behavior of sulphur in the basic open-hearth process has been delayed by lack of dependable data covering a wide range of slag conditions in the absence of other complicating factors that are likely to be present under operating conditions. It has been generally accepted that strongly basic slags will dissolve more sulphur than will those in which the acidic and basic constituents are more nearly balanced. It has been known that both slag and metal would absorb sulphur from the flame when the sulphur content of the fuel is high. It has been generally believed that manganese assists in the transfer of sulphur from metal to slag. Almost nothing has been known of the effect of temperature upon desulphurization; the data have been conflicting and there has been a tendency to depend rather heavily upon analogy with the better understood desulphurizing reaction of the blast furnace. Many attempts have been made to build up from open-hearth data a satisfactory theory of the chemical reactions involved in the process. That this has been an extremely difficult task is owing to two reasons that have been very generally recognized. The first is the parallel and simultaneous reaction of sulphur transfer between flame and bath. The second is the difficulty of distinguishing between the factors that ultimately limit the transfer from metal to slag-i.e., equilibrium conditions-and those whose influence is only upon the rate of transfer. It has been the object of the work here recounted to study the equilibrium distribution of sulphur between slag and metal in the absence of these complicating factors, with the hope that a better understanding of the equilibria involved would lead to a better control of the over-all process in the open-hearth furnace. A review of the literature on steelmaking reveals almost complete agreement with respect to the paramount importance of lime in promoting desulphurization. In expressing this effect in quantitative terms, however, and in formulating the relationships between the concentrations of other slag components and the desulphurizing power, there is little agreement. For instance, the "basicity" or basic strength of the slag is commonly measured by some kind of ratio of concentration of basic to acidic oxides. Of these, the following have been reviewed by Schenck :20 [ ] Among others there appear the following:

Jan 1, 1946

By Thomas. B. D.

GRAVITY concentration is a general term designating processes for separating and sorting granular material by means of forces that depend on the density, size and shape of the particles. When these forces are applied to the particles the latter are caused to move, those of a given density along one path and those of a different density along another, until ultimately they are sufficiently separated to be removed from the operation of the process. The forces producing this separation are in general complex and, unfortunately, no completely rigorous analysis of them has as yet been devised. The science of the hydrodynamics of viscous fluids deals with , the little that is known about the problem. Every observer of nature is familiar with the fact that an object moving in a fluid encounters a force of resistance. We have only to walk against a brisk wind or stir a cup of coffee, or observe the slow fall of a grain of sand in water, to experience directly the fact that such forces exist. The exchange of force between an object and a fluid in which it is moving operates on the fluid as well as on the object. Thus a sailboat may be set in motion by the force exerted by the wind on the sail; but the sail also deflects the wind-furthermore, the water in which the boat floats is set in motion by the force exerted by the boat as it cleaves through the water, The wind is thus enabled to do work on the water through the intermediary of the boat. In general, an object moving through a fluid encounters a force that opposes the motion and also acts to change the motion of the fluid. In gravity concentration we are dealing with this force of resistance acting on the granular particles that are to be separated. Furthermore, the motion induced in the fluid by a particle will in turn affect the motion of a neighboring particle. This effect will be reciprocated and in consequence there will be an exchange of forces between all the particles in the material. These effects are complex and the most that can be expected is that we may be able to determine in a qualitative way how they operate. In gravity concentration, then, the following types of forces may be distinguished as acting on the particles:

Jan 1, 1943

By A J. MacDonald, M P. Raffield

The South Deep massive orebody consists of 20 to 30 m of payable conglomerates ranging in dip from 11¦ to 20¦ and at depths of between 2400 m and 3400 m. The mine design process selected a mining method with a proven record of application that could be adapted to overcome the inherent difficulties associated with the depth, thickness and structural influences of the South Deep orebody. A modified cut and fill mining method provides a concentrated and highly flexible method of extraction. An integrated destressing cut will enable the modification of the stress environment such that it is suitable for massive mining operations. The choice of backfill type and stoping width in the destressing cut ultimately determines the characteristics of the stress modification and, as such, is critical to the success of the cut and fill operation. This paper discusses the decision process undertaken to determine the optimum backfill type and stoping width combination in terms of, conditions in the destressing cut faces and in the massive mining environment. Numerical modelling has been extensively used to analyse various scenarios to aid in the decision process.

Jan 1, 1998

By Terence P. Holman

Micropiles are employed to retrofit existing foundations, including underpinning prior to future construction operations and resupport of existing structures that have experienced distortion and settlement. This paper presents analyses of the two case histories requiring the use of micropiles in these applications. The first case history involves micropiles used to resupport a settled pump station next to an excavation which had suffered basal failure and ground loss. A compression load test pile experienced plunging failure which, after analysis of the strain-gauge data acquired during the test, was attributable to a greater degree of disturbance to the subsoils beneath the site than anticipated. After construction of the micropiles around the pump chamber a synchronized jacking system was used to lift and right the 320 kip chamber structure, followed by resupporting it on the newly constructed piles. The second case history entails the use of micropiles for direct underpinning support of existing columns prior to a utility tunnel installation within an existing high-bay warehouse structure. The design constraint that most significantly affected the successful completion of this project was the desire to bond the new piles into the existing reinforced concrete foundations. As a result of the coring method through the footing, the ultimate bond stress measured during a pullout test was more than twice that assumed in the design. Monitoring of the underpinned columns during trench excavation indicated that the system performance was acceptable. These case histories reinforce the concept that lower capacity micropiles can provide a rapid means to underpin and retrofit existing structures.

Jan 1, 2007

By S. Fox, Y. Fujita

Increasing demand for rare earth elements (REE) is expected to lead to new development and expansion in industries processing and or recycling REE. For some industrial operators, sending aqueous waste streams to a municipal wastewater treatment plant, or publicly owned treatment works (POTW), may be a cost effective disposal option. However, wastewaters that adversely affect the performance of biological wastewater treatment at the POTW will not be accepted. The objective of our research is to assess the effects of wastewaters that might be generated by new rare earth element (REE) beneficiation or recycling processes on biological wastewater treatment systems. We have been investigating the impact of yttrium and europium on the biological activity of activated sludge collected from an operating municipal wastewater treatment plant. We have also examined the effect of an organic complexant that is commonly used in REE extraction and separations; similar compounds may be a component of newly developed REE recycling processes. Our preliminary results indicate that in the presence of Eu, respiration rates for the activated sludge decrease relative to the no-Eu controls, at Eu concentrations ranging from <10 to 660 µM. Yttrium appears to inhibit respiration as well, although negative impacts have been observed only at the highest Y amendment level tested (660 µM). The organic complexant appears to have a negative impact on activated sludge activity as well, although results are variable. Ultimately the intent of this research is to help REE industries to develop environmentally friendly and economically sustainable beneficiation and recycling processes.

Jan 1, 2016

By P. A. Beck

ACCORDING to the traditional definition,1 re-x crystallization is a process taking place upon annealing of cold worked metals, characterized by the appearance of new strain-free grains, growing at the expense of the cold worked grains until ultimately the new crystals completely replace the cold worked structure. In metals like brass this process was observed long ago to take place simultaneously with the return of the mechanical properties to those characteristic of well annealed material. The above definition implicitly assumes that the matrix in which the new grains are growing is essentially still in the cold worked condition at the time it is absorbed by the new grains, i.e., that the energy stored in the metal during cold working is not greatly decreased by recovery, prior to recrystallization. Accordingly, the bulk of this energy is assumed to serve as the driving force for the process of recrystallization. Recently, a great deal of evidence accumulated, indicating that the traditional concept may be considerably in error. Suzuki&apos; found that by far the largest portion of the stored energy of cold work in compressed polycrystalline copper is released in the early stages of annealing, well in advance of recrystallization. Thus, the new recrystallized grains are undoubtedly growing in a matrix which is, at least energetically, no longer in the cold worked condition, but has nearly returned to the annealed state. It is true that in Suzuki&apos;s experiments the major hardness changes due to annealing were observed to take place concurrently with the formation and growth of the new grains, but this process apparently involved the release of only a minor residual fraction of the stored energy of cold work. Confirming in this respect the views of Wood," several recent investigations by various techniques lead to the conclusion that the grains of many, if not all, metals in the cold worked condition consist of small, slightly disoriented "crystallites" or "sub-grains".* (After appreciable deformation at room temperature, the subgrain size in copper appears to be about 7.104 cm and in high purity aluminum about 2.10-4 cm.) Recent investigations which demonstrated this phenomenon most convincingly are those of Heidenreich8 with hammered high purity aluminum, of P. B. Hirsch and collaborators7,8 with cold worked polycrystalline aluminum, copper, nickel, iron, gold, etc., and of Warren and Averbach9 with a rolled copper plus 2 pct Si single crystal. In addition to the existence of slightly disoriented sub-grains, the cumulative evidence from these investigations also shows that in the cold worked condition the subgrains themselves are severely strained. With regard to the early stages of annealing, Heiden-reich&apos;s results indicate that within the subarains of hammered high purity aluminum the strain gradually disappears, without discernible migration of the subboundaries. This recovery process presumably accounts for the major portion of the energy release

Jan 1, 1953

By Ellis Brien, O&apos

The fine coal fraction of a coal stream is that fraction which contains particles essentially in the -28 mesh size range. This fraction will normally first appear, in a coal preparation process, as a result of classification to segregate raw coal into various size components for beneficiation. More specifically, it could appear in desliming screen underfiow, classifying cyclone overflow, and in other process circuit components. It&apos;s recognized that the top size of this fraction has been determined more by secondary reasons, such as desliming screen area for the heavy medium cyclone circuit, the dewatering equipment for the fine clean coal, or even the stability of the refuse thickener circuit. Benefits, both economic and environmental, gained by beneficiating the fine coal fraction of a raw coal stream are increasing in number, until today, fine coal beneficiation is a necessity in most coal operations requiring any degree of preparation. This is due to some of the following reasons: • In the past it was frequently more economical to discard -28 mesh raw coal sizes than to recover the coal values from them. Today, mining costs have escalated to the point where all coal values leaving the mine should report to the finished product if possible. • Today&apos;s mining methods are producing run-of-mine coals containing higher percentages of -28 mesh sizes. This heavily weighs the economic balance in favor of beneficiating this fraction. As much as 25% of the run-of-mine coal from some mines is in the fine coal fraction. Beneficiation is recommended so that clean coal contained in the fraction can be recovered to reduce the loss of coal values. • Existing and pending environmental regulations mandate that solid fuels now consumed in power station boilers contain almost no pyritic sulfur. This form of sulfur appears in raw coal in the fine coal fraction, but can be removed by coal beneficiation. Environmental considerations dictate that every year, more of the fine coal produced for steam generating purposes in the US be beneficiated. To deliberately discard any portion of the fine coal can only add to this cost, not only in the sense that fewer tons will be recovered but, more importantly, additional tonnage will be diverted to the most capital- and labor-intensive section of the entire preparation plant-the fine refuse dewatering section. It is apparent to preparation engineers that dewatering and ultimate disposal of the fine refuse derived from this beneficiation step is certainly the most difficult technical problem presented in the entire plant design. Therefore, any beneficiation procedure which increases the fine coal recovery, or in any manner decreases the tonnage of fine refuse to be dewatered, is going to be cost effective. Since the -28 mesh fraction contains both coal and refuse particles in their finest particle sizes and thus presents the greatest degree of liberation, one would immediately infer that the highest degree of beneficiation would be achieved. This is not the case-the opposite is more likely true. Conventional coal cleaning devices based on specific gravity (such as

Jan 8, 1980

By W. H. Geis

The following pages summarize the causes that led to the suggestion of unit operation of the paloma field, the organization of com-mittees, preparation of the Paloma Operators Agreement and the reasons for rejection of the plan. Introduction Long ago an Austrian physicist observed that under certain conditions the condensation of a liquid from gaseous mixtures resulted from pressure reduction. Some 50 years later he would have known that his research was not in vain. He would have seen a group of serious operators and engineers seeking to apply that principle to a conservation . program for unitized production in the Paloma gas field—but he would also have been saddened to see that forces unsympathetic to the application of science in oil production, with little or no appreciation of modern methods and the advantages of cooperation, so far have prevented the adoption of that program. Soon after the Paloma field was discovered it was realized that all the hydrocarbons in the reservoir existed in the gaseous phase. This condition was confirmed over a large area. Therefore, if the field was to be properly exploited with the greatest ultimate recovery of condensate, the hydrocarbons in the reservoir must be produced in the single or gas phase. They must be cautiously extracted with the least possible reduction of the natural pressure. If this could be accomplished, or if artificial pressure control were utilized, the condensation and underground loss of the valuable distillate would be preventable—otherwise it would not be. Wholehearted cooperation of all oper-ators was required in order to accomplish the conservation purposes of the program. To this end the production engineers of the several companies concerned lent their time and talents. They fully investigated the physical relationships of the products within the complicated gaseous mixture, prepared extensive reports on the subject and drew up a reasonably efficient plan of operations, which now is commonly called the "Paloma Unit Plan" but actually was entitled "The Paloma Operators Agreement." All of the engineers were not in perfect agreement on all the features of the plan, or even on all the engineering principles. However, it is safe to say that on all the important or essential factors there was a unanimity of opinion, particularly as to the objectives. These were fairly simple and reduce to one controlling item, "pressure control" by cycling or recycling. Organization The principal operators initiated the program by calling a meeting of all the lessees in the area, at which an organization was perfected, known as the "Paloma Operators Committee." Various subcommittees were appointed to expedite the work, including an "Engineering Committee." These func-tioned in a businesslike manner, frequently reported to the superior committee and

Jan 1, 1942

By F. M. Lewis, J. F. Myers

The paper reports the development of a large, slow speed ball mill closed circuited with a hvdro-scillator. This increased grinding efficiency 28 pct over conventional units. AS the title indicates, this is a progress report for the first year of experimental operation of a relatively large diameter, slow speed ball mill with 1-in. balls, with and without various classifiers, and including a hydroscillator. Several years ago the authors came to the conclusion that if any further progress were to be made in our flotation process, an understanding of the chemical and physical reactions and the control thereof, in the grinding circuit, was essential. It may be stated, at this point, that the preparation of a correct feed for the flotation process is still the number one objective of our grinding studies. This progress report deals with grinding "per se," and not with the chemical and physical reactions as the result of grinding. Prior to this time we had taken the naive position, that since grinding had been practiced by the industry for some fifty years or more, it had been well exploited and developed by a host of investigators. Consequently, we had given the matter only superficial attention. At the start of our study, we were aware of the fact that some difference in opinion existed, but we were not prepared for the controversial evidence that developed nor for the extent to which the experts differed in their viewpoints. There seemed nothing else to do but to tour the continent and decide for ourselves what was right and what was wrong, read all the published literature and visit with anyone who had ideas about the matter. This we did over the next few ensuing years. Since our study has met with some measure of success, we report, herewith, these data that ultimately they may be of help to other investigators. In 1944 it became necessary to increase our grinding capacity and at the same time to grind the sulphide portion a little finer. This provided the impetus to put our ideas into effect. As has been previously described, our grinding flowsheet at that time consisted of a 6x12 rod mill, followed by two 5x10 ft, and one 6x12 ft trunnion overflow ball mills, with conventional rake classifiers. Naturally, we first considered all the possibilities of altering or adding to the existing equipment. It became evident that if any appreciable increase in grinding efficiency could be obtained, it would pay to scrap the existing small mills and classifiers and install one big mill with adequate classification. Trying to determine just how much more efficient a big mill could be, in which were incorporated all

Jan 1, 1951

By Suresh P. Babu, Joseph W. Leonard

The ultimate objectives of comminution should be to reduce material to some specified size while producing a minimum of undersize or fines, with less power, higher throughputs, and with more compact machinery. Attempts to achieve these objectives have been only partially successful, since gains in some areas are achieved at the cost of concessions in other areas. An example of some of these trade-offs can be found in the closed circuit crushing operations of the anthracite and crushed stone industry where, in order to produce a minimum of unprofitable fines, unfavorable concessions must be made in the size of the crushing facility where 2 or more crushing stages are interphased by 2 or more screening operations. This paper discusses experiments with a new process which attempts to achieve some, if not all, of the comminution objectives in a single pass operation without undesirable concessions; or, at the very least, to achieve more than is possible with systems in use today. These objectives are sought by use of water-flushing to enhance comminution. The crushing of coal and other minerals is a mature technology that is long overdue for a careful re-examination. Crushing of coal is generally achieved through the use of rotary breakers and roll crushers. Water-flushing, as envisioned here, appears to have potential application in the following areas: 1) For primary raw coal crushing or for secondary crushing of rotary breaker products. 2) For use as a size reduction and/or concentrating step for crushing coal mine refuse from present production or from existing refuse dumps. 3) For possible future application in underground preparation interphased with continuous mining machinery. Size reduction with conventional methods is known to be an inefficient operation when the increase in surface energy of the material crushed is compared with the input energy. A simple breakdown of the energy consumption in size reduction, given by Orr, is shown in [Fig. 1]. It is believed that most of the energy supplied is lost in the form of interparticle friction.2 Hence, the removal of fines immediately after their formation or the implementation of steps to prevent their formation, based on reduced friction, will reduce the energy requirements of the communition operation.3 A detailed discussion of the mechanism of size reduction is available in other literature.3-5 Even though the closed circuit crushing operation is designed to carry the fines away from the crushing zone, it is still inadequate for instantaneous removal of fines. The present investigation advocates the use of a flowing liquid medium to remove the crushed product immediately after size reduction. Unlike the wet-grinding methods, no constraints are required for the liquid flow rate, and, also, the liquid medium can be used in primary crushers like jaw crushers, hammer mills,

Jan 1, 1978

By Robert L. Hoss

The application of Muskat&apos;s differential equations for predicting the performance of a solution gas-drive reservoir to the Fullerton field indicates that a recovery may be expected of 14,030 bbl per acre under primary depletion, and 15,910 and I7,190 bbl per acre, respectively, under the two pressure-maintenance programs analyzed. In this analysis it is assumed that fluid properties are the same throughout the reservoir, that the core data used are typical of the Clear Fork limestone in the locality, and that injected gas is completely diffused through the pay zones. The economic study of the three recovery programs considered indicates that the one in which 60 pct of the produced gas is returned to the formation is the most profitable. Introduction The purpose of this paper is to consider the effect of pressure maintenance by gas injection on the ultimate recovery of oil from a solution gas-drive reservoir. This prediction of reservoir behavior is the result of calculations based on the equations presented by M. Muskat in the Journal of Applied Physics, March 1945. Reservoir The reservoir analyzed is the Fullerton field in Andrews County, West Texas. It is situated on a broad anticlinal structure with approximately 330 ft of closure, and extends approximately eight miles north and south and five miles east and west. The areal extent is taken as 16,642 acres. The producing horizon is the Lower Clear Fork limestone of Permian age, occurring at depths from 6650 to 7350 ft. It is divided into four zones. For the purposes of this study, the fluid characteristics in all zones are assumed to be the same. Therefore an equivalent zone is assumed of thickness equal to the sum of the thicknesses of the four zones, or an average of 216 ft. The reservoir contains 3,586,991 acre feet of producing limestone, with an initial volume of stock tank oil in place of 918,-403,100 bbl. The average connate water saturation is 24.0 pct. The permeability ranges from 6.7 md. to 12.6 md., and the porosity ranges from 7.7 pct to 11.4 pct, except for occasional cavernous streaks. Based on these values, the limestone may be considered intergranu1ar.l Production histories and material-balance studies have indicated that the driving mechanism is solution gas. Water levels have been established. Theory It is possible to express mathematically the change in residual oil saturation with pressure decline in a solution gas-drive reservoir as being equal to the change in the physical properties of the fluids present in the reservoir as this pressure decline occurs. Essentially these equations, in differential form as presented by M. Muskat,2 express quantitatively the instantaneous net effect

Jan 1, 1948

Brand se Baai, on the North-Western Cape Coastline of Southern Africa, is the site of the Namakwa Sands Limited Heavy Mineral Sands Mining Operation. The mine is a 14 000 000 tonnes per annum open-cast strip-style mining operation using truck and excavators to mine the sand. The orebody consists of relatively unconsolidated sands reworked by fluvial, marine and aeolian mechanisms. The total orebody includes three primary ore types, namely the Red Aeolian Sands, the Dorbank (a hardcap layer) and the Orange Feldspathic Sands, which forms the bulk of the resource. Mine planning and scheduling is done using a 50 m + 50 m variable bench height system with three primary bench heights. The Aeolian and Dorbank each have a bench that varies with their thickness of up to a maximum eight metres and the Orange feldspathic sand, a standard bench height of eight metres. For a number of years Namakwa Sands used a complex 2D, Datamine Based, macro system to perform the resource evaluation and mine design. As the mining operations expanded it became clear that a complete system redesign was required to cope with the increasing operational pressures coming to bear on the old system. The biggest challenges facing any replacement system are the sheer scale of the mining operation (16 km long and 7 km wide resulting in a minimum of 30 000 individual 50 m + 50 m mining blocks), the size of the geological data files (up to one gigabyte) and representing the benching system accurately (as the benches follow the topography and are not horizontal). In addition to these large-scale issues, the ability to run multiple mine design scenarios was required, an aspect which the old system was unable to perform adequately. Based upon these primary difficulties, a new business process was developed which required a complete rethink of the mine design methodology as well as the creation of a new system to deplete the resource model, produce monthly reconciliation evaluations and create a ultimate pit limit based on the operations mining cost model. Using a system based on the Earthworks Mine2-4D Mine Design and Scheduling System (with its mine design and scheduling specific tools), as well as a small, customised program, massive time-savings have been made. It is now possible to run several life of mine scenarios in the same timeframe that a single life of mine run took under the old system. This paper will present the methodology and business decisions behind the new Namakwa Sands system, and detail the restructuring and development used to deal with the large data set sizes and long processing times inherent in the old system. Future enhancements to the system that will make it more generically applicable to open-cast mining scenarios will also be dealt with.

Jan 1, 2003

By C. P. Proctor, E. D. Hummer, J. A. Younkins

COMPARATIVELY little has been published relating practical experience with the control of the solid content of washery water systems that must be "closed." A "closed" system is one that for some reason may not be cleaned by the emission of any effluent of high solid content to be replaced by clean water in order to hold the solid content of the waste water within reasonable bounds for efficient washing. It is the purpose of this paper to describe practical experience with such a water system and to present data that it is hoped may be of some use to those who are faced with a similar problem. The washery clarification system described comprises three thickeners and three filters, and although it is admitted that each of the fields of water settling and thickening and filtration offers a wide scope for investigation, the authors are not attempting to present the answers to these problems but to set down the results of some experience along these lines, making no claims as to the ultimate solution of the problems in hand. The subject of water clarification has vast ramifications; for example, the science of flocculation as applied to water settling and filtering opens up fields of wide possibilities of which as yet little is known. A comparatively recent entry into the field of water treatment, which may offer possibilities of practical application, is the use of "wetting" agents such as "Aerosol." GENERAL ARRANGEMENT OF WATER SYSTEM So that the reader may obtain a comprehensive picture of the factors concerned in the control of a closed water system, it is necessary to give a brief description of the washing plant that is serviced by the water system to be described in the following pages. As we are not primarily concerned with the washing plant in detail, only a brief description of the flow of coal as it relates to the water system will be given here. For a more detailed account of this type of washer, the reader is referred to earlier papers.1.2 The washer, which handles 0 to 4-in. raw coal (750 tons per hour) is an American Rheolaveur plant of two complements; the coarse coal or sealed discharge plant and the fine coal or free discharge plant. The sealed discharge plant (Fig. r) consists of twin 48-in. primary launders, each having two washing boxes, supplemented by one 32-in. rewash launder, which also has two washing boxes. "Push" water for washing and conveying coal in these launders is supplied from a constant-head tank under pressure head of 10 ½ ft. at the point of introduction at the feed end of the launder. This tank also supplies water for upward or vertical currents used in the Rheolaveur or washing boxes. Each box discharges into a sealed boot elevator (Fig. I).

Jan 1, 1942