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By R. F. Krueger

Results are presented from tests of dynamic fluid-loss rates to cores from clay-gel water-base drilling fluids containing different commercial fluid-loss control agents (CMC, polyacrylate or smt,ch), organic viscosity reducers (quebracho and complex metal lignosulfonate) and oil at several different levels of concentration. In the dynamic system the most effective individual additives to the clay-gel drilling fluid, based on cost-equalized concentrutiom, were found to be starch and the viscosity reducers. These results do not conform with the rankings determined by API fluid-loss rests, which indicate CMC, polyacrylate and starch to be the most effective and comparable. Generally, minimum dynamic fluid-losr rates were attained at cost-equalized concentrations of additive (including thinner) of about $1.00/ bbl, or less. For chernically treated clay-gel drilling fluids, both the standard and the high-pressure API filter-loss tests were found to he inaccurate indicators of trends in dynamic fluid-loss rates under the test conditions used, particulurly for drilling muds containing viscosity reducers. From a practical field viewpoint, restrictions on the applicability of the API fluid-loss test are such that it is open to question whether or not results of this test can be used routinely with confidence as an indicator of control of down-hole fluid loss under field treating conditions. INTRODUCTION The petroleum industry spends large sums of money during drilling operations to control the fluid-loss properties of drilling fluids based on the standard API filter-loss test,' which is a static filtration system. Laboratory studies' ' of dynamic filtration have shown that in a given time period filtrate loss from a circulating mud stream is greater than from a static system and that it is a function of linear mud velocity, pressure and the properties of the drilling fluid. Ferguson and Klotz' and Horner, et al," observed that (I) the dynamic fluid-loss rates for the drilling fluids used were not related to the extrapolated API filter loss and (2) the drilling fluids with the lowest API filter losses did not have the lowest dynamic fluid-loss rates. However, there has been no published information on the relative effects on dynamic fluid-loss rate as a given drilling fluid is treated with increasing amounts of chemical additive to reduce the API filter loss. Such information is economically important because drilling-fluid costs rise rapidly as chemical requirements increase. This paper presents the results of a study of dynamic filtratioi rates to cores from a clay-gel water-base drilling fluid treated with various commercial viscosity reducers and chemical fluid-loss control agents. The dynamic fluid-. loss rates to cores are compared with the standard API filter-loss values at several different levels of additive concentration. Dynamic filtration rates were obtained in each experiment under two different simulated wellbore conditions: (1) filtration just above the bit through a new mud cake laid down dynamically on a freshly drilled formation and (2) filtration up-hole through a mud cake formed by deposition of a static filter cake on top of the initial dynamically formed cake. The latter case corresponds to the bottom-hole conditions existing above the bit when mud circulation is restarted after a stand of pipe has been added or a round trip has been made to change the bit. Except for the short-duration, high-rate filtration beneath the bit where no mud cake can form, these two conditions probably represent the two extremes of dynamic filtration. Because thickness of a dynamic mud cake formed on freshly exposed formation is limited by the shearing action of the mud stream, the filtration rate for this condition is high. On the other hand, once circulation is stopped and a static mud cake forms on top of the dynamic cake, re-starting circulation has only a small effect on the cake properties and filtration rate is much lower thereafter. A discussion of the mechanics of mud-cake deposition and dynamic filtration is outside the scope of this paper but may be found in more detail in publications by prior investigators. APPARATUS AND EXPERIMENTAL CONDITIONS The test equipment used to simulate the dynamic flow conditions existing during drilling was a modification of that described previously by Krueger and Vogel: A schematic flow diagram is shown in Fig. 1. In general, a power-driven, high-pressure mud pump capable of delivering up to 60 gallmin was used to circulate drilling fluid parallel to the faces of 1-in. diameter sandstone cores mounted in a 2 3/4-in. ID high-pressure test cell. Pump rates were controlled by means of a magnetic clutch to maintain an average axial fluid velocity of 110 ft/min in the annular space between the cell wall and a 1 1/2-in. rod positioned on the center line of the cell. The core specimens were Berea sandstone plugs sealed with plastic inside 1 1/8-in. OD tubes and were fluid-saturated prior to use. Burettes were used to accumulate fluid discharged from the cores. The mud sump shown was used for treatment and storage of the drilling-fluid samples during a particular test. The valve arrangement permitted either (1) circulating drilling fluid through the by-pass line while treating with

By Yang Ge, Yuzhu Pan, Mansheng Chu, Jue Tang, Cong Feng, Lihua Gao, Zhenggen Liu

The low-grade nickel laterite ore used in this study was successfully upgraded through selective reduction by industrialized rotary kilns followed by magnetic separation. The results indicated that both iron metallization extent and nickel metallization extent showed obvious linear correlation. In addition, rational roasting temperature and prolonged grinding time could enhance the overall fineness level of reduced laterite nickel ore. The moderate roasting temperature at a range of 1100–1150 °C should be suitable enough to a conversion of total nickel to metallic nickel and partly inhibited the conversion of total iron. Ultimately, the obtained ferronickel alloy with a nickel content of 7.32% and an iron content of 84.15% was prepared successfully from the lowgrade nickel laterite ore containing a nickel content of 1.92%. Nickel and iron recoveries were 96.72 and 89.08% respectively, which achieved the highly efficient recovery and utilization of iron and nickel from low-grade laterite nickel ore

By P. H. Delano, Will H. Coghill, G. D. Coe

Data concerning the actual net energy required for pulverizing coal are lacking from the literature on coal pulverization. Power data given in the literature concern gross power and frequently include the power expended on the air blast or product-conveyor system, as well as power lost through friction in the pulverizer itself. Consequently, such data supply no information regarding the net energy consumed in pulverizing or the relative grinding resistance of various coals. In the study of coal pulverizing much work has been done on what is known as " grindability-index numbers," but grind-ability tests have seldom considered the energy input (that is, the "cause" part of the system "cause and effect"). There are many records of the effect, but even these have questionable value because the "grindability" tests did not make a pull'erized product by a method comparabLe with- commercial pulverizers. In many of the tests, only a selected part of the coal was ground, and the remainder continued to be an unknown factor. Furthermore, none of the grindability tests provide the balanced circulating load common to all modern commercial pulverizing. In order to obtain better understanding of the fundamental principles of coal pulvcrizing, the Bureau of Mines is investigating the "cause" phase; that is, the net energy input or net power excluding friction between the moving parts of the machine. As a first step in the investigation,it was necessary to design and construct a laboratory machine in which no friction-loss corrections are involved, which could grind and finish a minus-¾-inch feed, which could be emptied rapidly and completely, which could operate with a continuous feed and finish in an air separator; and, most essential of all, which would be equipped with a measuring device for recording only that portion of the total power consumed in actual pulverizing, hereafter referred to as "net energy input" or "net horsepower-hours." The channel-roller pulverizer hereinafter described is believed to meet these requirements. History oF the Machine The, work started when Ralph W. smith, engineer, St. Genevieve Lime Co., St. Cenevieve, Mo., was appointed one of the consultants of the Bureau of Mines, United States Department of the Interior, in cooperation with the University of Alabama, to develop a machine for testing the abrasiveness of coal. Smith had already developed a machine for testing the abrasive-ness of dentifrices,1 and it was thought the same principles might be applied to coal, Smith's machine had a revolving disk with an annular channel in which the sample was placed. Riding in the channel was a small block or runner of selected soft metal held

Jan 1, 1942

By Will H. Coghill, P. H. Delano, G. D. Coe

Data concerning the actual net energy required for pulverizing coal are lacking from the literature on coal pulverization. Power data given in the literature concern gross power and frequently include the power expended on the air blast or product-conveyor system, as well as power lost through friction in the pulverizer itself. Consequently, such data supply no information regarding the net energy consumed in pulverizing or the relative grinding resistance of various coals. In the study of coal pulverizing much work has been done on what is known as " grindability-index numbers," but grind-ability tests have seldom considered the energy input (that is, the "cause" part of the system "cause and effect"). There are many records of the effect, but even these have questionable value because the "grindability" tests did not make a pull'erized product by a method comparabLe with- commercial pulverizers. In many of the tests, only a selected part of the coal was ground, and the remainder continued to be an unknown factor. Furthermore, none of the grindability tests provide the balanced circulating load common to all modern commercial pulverizing. In order to obtain better understanding of the fundamental principles of coal pulvcrizing, the Bureau of Mines is investigating the "cause" phase; that is, the net energy input or net power excluding friction between the moving parts of the machine. As a first step in the investigation,it was necessary to design and construct a laboratory machine in which no friction-loss corrections are involved, which could grind and finish a minus-¾-inch feed, which could be emptied rapidly and completely, which could operate with a continuous feed and finish in an air separator; and, most essential of all, which would be equipped with a measuring device for recording only that portion of the total power consumed in actual pulverizing, hereafter referred to as "net energy input" or "net horsepower-hours." The channel-roller pulverizer hereinafter described is believed to meet these requirements. History oF the Machine The, work started when Ralph W. smith, engineer, St. Genevieve Lime Co., St. Cenevieve, Mo., was appointed one of the consultants of the Bureau of Mines, United States Department of the Interior, in cooperation with the University of Alabama, to develop a machine for testing the abrasiveness of coal. Smith had already developed a machine for testing the abrasive-ness of dentifrices,1 and it was thought the same principles might be applied to coal, Smith's machine had a revolving disk with an annular channel in which the sample was placed. Riding in the channel was a small block or runner of selected soft metal held

Jan 1, 1942

By C. Pin, J. Bruce Wagner

In order to demonstrate the role of defect chemistry in the solid state to a process-metallurgy reaction, laboratory experiments were carried aut to remove copper from liquid lead using lead sulfide which had been doped to provide definite concentrations of cation vacancies. The removal of copper decreases in the order: bismuth- or antimony-doped PbS, undoped PbS or tin-doped PbS, and silver-doped PbS. The data are explained using the simple mass -action law and in terms of the exceedingly rapid interstitial diffusion of coppw in PbS to vacant cation sites. Results are in agree-ment with thermodynamic predictions and certain industrial practices that copper may be removed from lead below the limit of its coexistence with pure Cu,S and pure PbS. The role of defects in solids has been exploited in semiconductors and many device applications, in explaining tarnish formation, dissolution rates, and mechanical properties. The application of these principles to many process-metallurgical problems has not received as much attention. It was the purpose of this study to illustrate the role of defect structures of solids in a process-metallurgical application. When lead is refined, the ore, galena (PbS), is reduced by a series of operations to liquid lead (lead bullion) containing impurities such as copper, bismuth, antimony, arsenic, and silver depending on the original composition, the fluxes added, and the particular type of operation, see Fig. 1. In order to produce a pure lead, a series of refining operations are carried out. A typical lead-bullion analysis before drossing is shown in Table I. Copper is removed first as its presence increases the zinc consumption during the desilverizing process. For preliminary copper removal the lead bullion is brought to a temperature of 440°C in a kettle of 60 metric tons capacity. Some tar (or sawdust or resins) is usually added and the kettle is cooled slowly to 400°C while the liquid is stirred. A solid phase, called a dross, floats on top of the liquid and the particles of the solid phase adhere to the tar. This "hot dross" is collected. It contains mostly lead and copper which are partly oxidized and/or sulfurized if the lead bullion contains some sulfur.*

Jan 1, 1963

THE fall meeting of the Petroleum Division at Fort Worth, Texas, Oct. 19 and 20, attracted even a larger number of members and guests than had been expected. Over three hundred registered the first day though the attendance and interest in the program was so keen that it was necessary to begin technical sessions without. waiting for registration. Members were present from as far away as Venezuela and while the larger attendance was from the Mid-Continent and Texas fields, California and the eastern states were represented by influential delegations. One pleasing feature of the meeting was the presence of a number of the older executives of the oil industry who evi-dently have come to realize the importance of the work the Division is doing in advancing technology. The meeting was held under the auspices of the Production Engineering section of the Division. E. 0. Bennett, chief production engineer for the Marland Oil Co. of Texas, and chairman of the Fort Worth Sec-tion. of the A.I.M.E., served as general chairman. J. B. Umpleby, president of the Petroleum Reclamation Co. and vice-chairman of the. Division was in charge of. the program with a special committee in charge of the technical sessions, of which A. C. Rubel of the Union Oil Co. was chairman. The sessions were held in the ball room of the Texas Hotel, which also served as general headquarters. The Wednesday morning session was devoted to the air-gas lift, John M. Lovejoy, chairman of the Division, presiding. The first paper, on the "Principles Under-lying the Air-Gas Lift." by J. 0. Lewis, was presented by R. R. Brandenthaler and discussed by Messrs. Love-joy and Millikan. Mr. Brandenthaler's own paper on the "Effect of Gas Lift on Physical Properties of Oil" followed and was discussed by Messrs. Umpleby, Walker, Mills, Bennett, Williams and Van der Gracht, the dis-cussion turning mainly on what is known and what unknown regarding oxidation of oil under various con-ditions. The third paper of the morning, "Handling Recirculated Gas," by R. D. Gibbs, was presented by Mr. Taylor. The discussion that followed was general and brought out a large amount of data on the costs and results of compression versus, absorption in total gasoline recovery from gas. The next group of papers dealt with "Mechanical Installations of Air-Gas Lift," the first by E. V. Foran on the Texas district outside the Gulf Coast, the second by L. L. Brundred on the Gulf Coast, the third on Kansas and Oklahoma other than Seminole by R. W. Bond, the fourth on Seminole by C. R. Swartz, and the fifth relating to California by H. C. Miller. Some of these were read by title in order to give more time for discussion which was active and participated in by. Mills, Griswold and others.

Jan 11, 1927

By W. H. Meyer

Although an aerial photograph is not a map, most of the information that is necessary for compiling a map is recorded in the photograph provided some form of radial-line method is used to determine the true position of images. The slotted- templet method, which is based on the well-established radial-line principle, has made it possible to construct accurate mosaic maps at a cost comparable with that formerly charged for the average mosaic. It makes possible the securing of these results with far less ground control, which is always expensive. It has expanded widely the possible use of aerial photographs for mapping. ALTHOUGH a vertical or oblique aerial photograph is not a map, most of the information that is necessary for compiling a map is recorded in the photograph. Some ground controls or data are needed, the number and accuracy to be governed by the mapping method to be used and the scale and final accuracy desired in the finished map. RADIAL-LINE CONTROL It is essential that the principle of radial-line control in the use of vertical aerial photographs be understood and accepted before any confidence in the slotted-templet method can be gained. A simple sketch and some enlargements showing the Empire State Building will serve to illustrate the cause of displacement of an image that is above the datum plane as recorded in the photograph. The top of the tower of the Empire State Building is directly above the base, and when the image of the top is shown almost two blocks away from the center of the base, it is evident that the image of the top of the tower is not the true location, and in this particular case we know where the true location is; namely, over the center of the base of the tower. If the top of the tower were the top of a hill 1250 ft. high, we would have no way of knowing by inspection the true position of the top of the hill, so it is very important to have a method whereby the true location of any such point can be determined, and thereby make it possible to convert the detail of an aerial photograph into accurate information for mapmaking.

Jan 1, 1939

By D. E. Hirsch, J. W. Davis, Norman G. Foster, F. G. Schwartz

This investigation was undertaken to identify nitrogen , oxygen , and sulfur compounds (polar components ) found in catalytically cracked distillate fuel . It is significant that these compounds concentrate in the sediments and gum that form when the fuel is stored . The polar components were divided in three portions by elution from a chromatographic column filled with alumina . After the hydrocarbons were washed through the column with pentane , a benzene eluted extract was removed followed by an ether eluted extract and then an alcohol desorbed fraction . These extracts were individually distilled in a molecular still and the fractions were analyzed by gas chromatography , mass spectrometry , infrared spectrometry , and elemental and functional analyses . Naphthalene and its homologs , phenanthrene and its homologs , and other compounds such as methyl indene , C₂ quinolines , phenyl indan , and the benzothiophenes , appear in the benzene extracts . Quinolines are the major components of the lower boiling fraction of the ether extract , although carbazoles , acridans , and indoles are present in some concentration . Elemental and functional group analyses of the alcohol fraction indicated 10 percent of the distillate portion was nonhydrocarbon , and in this portion the average molecule contained more than one nonhydrocarbon element . Compounds of such complexity are not well known , and standards for positive identification are limited .

Jan 1, 1963

By W. G. Stevenson

With the author's background of extensive experience, this paper delineates the role of the prospector in modern-day exploration programs. The stages in the development of a prospect to a producing mineral district are outlined, and examples are given from the Ross River area of the Yukon Territory and the Peachland and Highland Valley areas of British Columbia.

Jan 1, 1970

By Raj K. Rajamani, Sanjeeva Latchireddi, Dave Plummer, Stieger. Julius

"The Cortez SAG mill liner design was recently modified to improve the mill operating energy efficiency. In addition to the expected benefits, the design change unexpectedly played a significant role of verifying the proposed grinding circuit expansion for the Cortez Hills project as well as illustrating current SAG mill bottlenecks. This paper will discuss the evaluation that addressed operator and maintenance observations, historical PLC data, and “crash stop” observations to calculate charge motion, packing between lifters, and slurry transport through grate and pulp lifters. The performance of the new shell and discharge end liner design will also be discussed as well as their role in the possible grinding circuit expansion.INTRODUCTIONCortez Gold Mines commenced operation in 1969 with its Mill #1 rated at a nominal 2,000 tons per day. Mill #1 was a conventional rod mill/ball mill circuit followed by an 8-stage CIP circuit. In the early 1990s the Pipeline deposit was discovered and subsequently Mill #2 was built to process the increased ore volume. Mill #2, commissioned by mid-1997, was a nominal 10,000 tons per day SABC circuit followed by an 8-stage CIL circuit for oxide gold ore. Operating and commissioning details of this plant have been published elsewhere (Powell, 1998). By late 1998, there was not enough ore volume to support both mills so the smaller, higher operating cost Mill #1 was put under its current care-and-maintenance status. A simplified flowsheet of the current mill is shown in Figure 1."

Jan 1, 2007

By E. D. Gardner

About five minutes after 3 o'clock on Thanksgiving morning, November 24, 1927, smoke was discovered coming up No. 2 shaft of the Magma Copper Co.'s mine at Superior, Pinal County, Arizona. Of the 49 men in the mine at the time, 42 were able to make their way to the outside unassisted, but 7 perished. Six bodies were recovered, and the seventh, that of an Apache Indian cage-tender at No. 2 shaft, is thought to be in the caved section between the 1,600 and 2,250 foot levels A second and independent fire broke, out in No. 1 shift about 2 a.m., November 27, burning .out the shaft timbers from the surface to the 200-foot level and destroying the surface fan and structures. About 20 men were underground fighting the fire in No. 2 shaft, but all were safely withdrawn, aril the fire in No. 1 haft was quickly controlled.

Jan 1, 1928

By M. C. Malamphy

A general description of the activities and results of bauxite investigation by the Bureau of Mines at its bauxite project in Saline and Pulaski County, Ark., has been given in Report of Investigations 4251. The factual data obtained during the project are reported in a series of Reports of Investigations beginning with No. 4251, of which this is the eighth. The data are accompanied by maps showing the location of holes drilled and the deposits developed. Each area upon which work was done is reported by half townships. Each bauxite deposit investigated was assigned a name and a number for ready reference and identification, and its location and name are shown on the appropriate half-township map. The description of each half township is no followed by a description of each deposit developed in that half township. The 14 half township maps and 77 detail maps of the bauxite deposits in this series of reports give the locations and numbers of all the holes drilled on the project, except the deep holes drilled to test the geophysical anomalies. The locations at those holes show on figure 13 in volume I. The descriptions of the deposits developed are followed by complete detailed logs at the drill holes. Many at the widely spaced exploratory holes did not discover bauxite of minable thickness and gra.de. The pertinent information obtained from the widely spaced exploratory drill holes has been summarized and tabulated, and a table for each halt township is given with the discussion covering that area.

Jan 1, 1948

By Peter Altounyan

The rate of decay of the energy from a hammer blow to a coal mine roof is related to the size and nature of the contacts between the rock being struck and its surroundings The use of this principle as a basis for testing roof condition in South African coal mines has been investigated under a SIMRAC funded project, with the objective of reducing the number of roof falls. ( I ) A prototype instrument having a simple numeric display related to the energy decay rate (the Acoustic Energy Meter) was constructed and used to carry out surveys at mines covering the range of coal mine roof geology and condition. The equipment was simple to use and provided excellent discrimination between visually intact roof and less stable conditions, once the normal/abnormal threshold had been established for a given site. Areas of potential roof failure were identified close to roof slips and igneous dykes, and confirmed in areas of visible deterioration. Directional stress effects on roof condition were also identified at some sites, with consistent differences in readings obtained depending on the direction of drivage. Repeat surveys also indicated areas where roof conditions were deteriorating with time. Based on the success of the prototype, a hand held version of the instrument. has been developed to a specification from Anglo Coal. This incorporates a simple 'traffic light' green / amber/ red display, making the device suitable for routine use by mine operatives This version is now undergoing trials at an Anglo Coal mine. It is concluded that the Acoustic Energy Meter shows great potential as a test instrument providing output which relates to the stability of the immediate roof.

Jan 1, 2000

By Werner Schwartz, Wolfgang Haase

Lead sulfide concentrates, which may include other lead concentrates, are sintered on an up-draught sintering machine without the addition of any diluting agents or fluxes. Subsequently they are melted in an oil- or gas-fired rotary furnace. The sintering and melting processes are based upon the following roast-reaction: PbS + 2 PbO = 3 Pb + SO, PbS + PbSO, =2 Pb + 2 SO, For obtaining a lead bullion free from sulfur, the sintering process is carried out in such a way that the sinter product contains a small amount of excess oxygen above that to react with the sulfides. At the end of the melting process, when the reactions are finished, the remaining small amount of oxide residues is reduced with coal to which a certain percentage of soda ash (about 1 pct of the lead bullion) is added. For the lead smelting process described neither coke nor fluxes—except soda ash—are required. This process is being utilized by a European smelter successfully and with a high lead recovery. The consumption figures for the smelting of 100 tons per day of lead concentrates are indicated. The lead content of the lead concentrates from modern ore dressing plants ranges from 65 pct to above 80 pct. In most lead smelters of the world these concentrates are smelted in a blast furnace. For blast-furnace smelting the concentrates have to be desulfurized and agglomerated by sintering. A requirement for the perfect operation of a down-draught sintering machine and of a blast furnace is a maximum lead content in the feed of 40 to 45 pct. For this reason, some lead concentrates have to be diluted by adding return slags, limestone, and possibly iron oxide and sand. As an example, 100 tons of lead concentrate with 72 pct Pb would contain 13.5 tons of gangue (including the zinc). To produce a perfect sinter with 42 pct Pb it would be necessary to add 70 tons of flux and return slag, more than five times the original weight of the gangue, to the sinter mix and blast-furnace charge. A correspondingly large amount of coke would be required in order that all of these materials reach the heat of formation and the melting temperatures of the slag (1200" to 1400°C) inside the blast furnace. The roast-reaction process presents a possibility for lead recovery without dilution of the concentrates. In this process the concentrate mixed with coal is placed upon a Newnam-hearth and air is blown through nozzles into the heated mix. AS a result metalllic lead and a relatively great amount of so-called .'Grey Slag" with a lead content of 25 to 35 pct are formed. The slag is sintered to eliminate sulfur and, after addition of the requisite fluxes, treatt:d in a blast furnace. Owing to the poor recovery of lead from the hearths and to the unavoidable heavy hand-work plus the risk of poisoning this process is utilized in very few 112ad smelters today. Since in mxny countries of the world coke is expensive and difficult to obtain, it appeared feasible to use the principle of the roast-reaction by modern sintering and melting methods with recovery of the lead in electric, or oil, gas, or coal-fired furnaces. Two processes are utilized on an industrial scale: A) Lead smelting in the electric furnace of the Bolidens Gruv A/B in Sweden, as described by S. J. Walldcn, N. E. Lindvall, K.G. Gorling, and S. Lundquist. B) The self-fluxing lead smelting of Lurgi Gesell-schaft fiir Chemie und Huttenwesen m.b. H., Frankfurt a M, Germany, which is described in this paper. In the Boliden process referred to above the sinter mix is pelletized by enveloping return fines with layers of flue dust, limestone powder, and dried galena concentrate. The roasting and agglomeration are carried out on a down-draught machine, and a slight excess of sulfur is left in the sinter product. During the smelting in the electric furnance the roast-reactions occur and a slag poor in lead and a sulfur bearing lead are formed. This latter is subsequently oxidized in a converter to obtain lead bullion and dross. The Lurgi-process achieves the maximum possible extent of the roasting reaction on the sintering machine. The wet flotation concentrates are blended with return fines (lead content 70 to 80 pet), any existing flue dusts and lead slimes—but without the

Jan 1, 1962

By Ing. Ludwig Schlebusch

The significant growth of the world population and its demand for an increasing standard of living leaves no other choice than continuously augmenting industrial production and making available steadily greater quantities of raw materials at still better qualities. Therefore, any raw material has to be utilized to the greatest possible extent and processed at optimum economy on large mechanical equipment. The layout of processing facilities of large throughput rates can be based on multiple or single lines. The latter enables the application of large machines of optimum size. Proven equipment for homogenizing, classifying, separating superfine material, sorting and dewatering are-now available in optimum sizes. By the use of these large machines operating cost of pit coal preparation plant can be significantly lowered by reducing the specific requirements regarding space, capital investment, energy, operation and maintenance.

Jan 1, 1980

By Stanley C. Suboleski

1977 was the modern day low point for productivity in the coal industry. At that time, no one could have predicted that the industry was about to enter a productivity rise that would continue unabated through the 1980s. Thirteen years later, the cause of this phenomenon is still not well understood. While many government studies focused on the fall in productivity from 1969 to 1977, none have been undertaken to explain the rise. Some of the likely causes of the gain in productivity during this period will be discussed here. An attempt will also be made to quantify the value of some of the technical innovations that have taken place. The remainder of the gain is assumed to come from non-technical sources. The nature of these sources, although largely non-quantifiable, is also discussed. Since the number of mines that have adopted the technical innovations noted here is largely unknown, the actual contribu¬tion of technological changes to the overall productivity gain can only be estimated.

Jan 1, 1990

During the past decade, since the appearance of Popoff's (1966) review of grade tonnage calculations, it has become evident that defining a new method which could be followed step by step by a technician and which would improve reserve estimation was not a simple thing to do. Some concepts have emerged though, and most definitions of ore reserve now require a statement of reliability (USGS, 1975; and Zwarten-dyke, 1976). Exactly how the reliability should be computed is not a straightforward process. It also became apparent that not only was reliability a requirement but that additional questions such as how to predict mining re- serves from geological reserves were to be asked. In fact what became obvious is that specific models have to be designed for each type of mine and for each type of question which is asked. Most methods can be encased within a unique general geostatistical, model, but it takes considerable custom tailoring, and there are still questions that are left unanswered. In this chapter, before giving techniques, we intend to first review what should be required from a grade tonnage calculation done at the exploration stage. If the final objective is still to specify a tonnage and grade, with eventually a reliability, many qualifications have to be made before quoting, and of course, calculating these two numbers. Then the currently avail- able methods will be reviewed and future trends will be shown; finally a few case studies will be given and recommendations about the choice of a model will be made.

Jan 1, 1979

Jan 1, 2000

By G. J. Slavens

The Bureau of Mines has devised technology to recover cobalt, nickel, and byproduct copper from domestic lateritic material using an oxidative, ammoniacal ammonium sulfate leach. Nickel, cobalt, and copper were recovered by solvent extraction and electrowinning. To reduce the cost and complexity of cobalt recovery, an alternate method using lig-nite to extract Co3+ was investigated as reported herein. In the first 25 BV (bed volumes) of solution contacted with lignite, 97 wt pct of the Co3+ was extracted in & downflow column at pH 9.3. After 85 BV, the lignite was loaded to 1.5 wt pct Co3+ and further ex-traction was minimal. The lignite extracted cobaltic ammine complexes rather than cobalt ions and was nonselective as to other -metal cations and ammonium ions. Deleterious cations and 80 wt pct of the nonammine ammonium ions were removed prior to acid elution of the lignite. The removal of the ammines prior to elution to produce an ammonium-ion-freeCo3+ eluate was only partially successful. Thus Co3+ removal from ammoniacal solutions by lignite is technically difficult and economically unfeasible. It was also determined that reduction of Co3+ to Co2+ occurred during the extraction-elution process.

Jan 1, 1984

By Ma Fangyuan, Li Min, Xu Bing, Xian Yushuai, Zhao Tonglin

The study explored a new combination amine collector (XK-28) that can obtain better flotation index and defoaming rate for magnetic iron ore. A pure quartz mineral test indicates that dodecylamine, XK-I, or XK-II has good collection effect on quartz and their best flotation pH is in the range of 6–7. The experimental results of a combination collector (XK-I:XK-II = 2:8.) show that the separation effect for quartz and magnetite is better than that of a single collector. The single factor for actual mineral test shows that when the combined reagent is used as the collector, the best conditions are pH = 7, 90 g/t dosage, and XK-I:XK-II = 2:8, which is consistant with the pure mineral test. The difference for dodecylamine as a collector is that the optimal dodecylamine dosage is 60 g/t. Under the best conditions of actual ore test, dodecylamine and XK-28 are respectively used as the collector to compare the flotation indexes. Study results indicate the magnetite grade using XK-28 as the collector is higher by 3.02% than that of using dodecylamine, recovery is increased by 1.92% over that of using dodecylamine. Based on the measurement results of zeta potential, uses of XK-28 can slightly increase the zeta potential difference of quartz and magnetite to enhance the flotation effect. The mechanism of action between the XK-28 collector and the actual ore from the Gongchangling plant is physical adsorption, and this adsorption also has hydrogen bond formation. Contact angle results shows that XK-28 improves the hydrophobicity of quartz better than dodecylamine. Compared with dodecylamine, XK-28 is more sparsely arranged on the surface of the foam. The foam produced by the XK-28 collector has better air permeability and it is more likely to merge and disappear with bubbles.

Jul 21, 2021