Search Documents

By William H. King

Preliminary examination of the Virgin River manganese deposit was made by C. H. Johnson, Bureau of Mines engineer, in March 1941. During, April and May 8 trenches were dug, a 48-foot inclined shaft was sunk, and 719 linear feet of channel samples were cut. McKelvey and Wiese, of the Federal Geological Survey, completed mapping the deposit, including the entire area of overlying formations comprising the point of land between the outcrop and Lake Mead. The geologic maps, maps, and ore estimates resulting from these examinations formed the basis for a preliminary report by C. H. Johnson, who recommended further exploration by diamond drilling.

Jan 1, 1949

By John J. Sammarco, Edward F. Fries

This report (Independent Functional Safety Assessment 4.0) is the fifth in a nine-part series of recommendations addressing the functional safety of processor-controlled mining equipment. It is part of a risk-based system safety process encompassing hardware, software, humans, and the operating environment for the equipment’s life cycle. Figure 1 shows a safety framework containing these recommendations. The reports in this series address the various life cycle stages of inception, design, approval and certification, commissioning, operation, maintenance, and decommissioning. These recommendations were developed as a joint project between the National Institute for Occupational Safety and Health and the Mine Safety and Health Administration. They are intended for use by mining companies, original equipment manufacturers, and aftermarket suppliers to these mining companies. Users of these reports are expected to consider the set in total during the design cycle. • 1.0 Safety Introduction.—This is an introductory report for the general mining industry. It provides basic system/software safety concepts, discusses the need for mining to address the functional safety of programmable electronics (PE), and includes the benefits of implementing system/software safety program. • 2.1 System Safety and 2.2 Software Safety.—These reports draw heavily from International Electrotechnical Commission (IEC) standard IEC 61508 [IEC 1998a,b,c,d,e,f,g] and other standards. The scope is “surface and underground safety-related mining systems employing embedded, networked, and nonnetworked programmable electronics.” System safety seeks to design safety into all phases of the entire system. Software is a subsystem; thus, software safety is a part of the system’s safety. • 3.0 Safety File.—This report contains the documentation that demonstrates the level of safety built into the system and identifies limitations for the system’s use and operation. In essence, it is a “proof of safety” that the system and its operation meets the appropriate level of safety for the intended application. It starts from the beginning of the design, is maintained during the full lifecycle of the system, and provides administrative support for the safety program of the full system. • 4.0 Safety Assessment.—The independent assessment of the safety file is addressed. It establishes consistent methods to determine the completeness and suitability of safety evidence and justifications. This assessment could be conducted by an independent third party. • 5.0 Safety Framework Guidance.—It is intended to supplement the safety framework reports with guidance providing users with additional information. The purpose is to assist users in applying the concepts presented. In other words, the safety framework is what needs to be done and the guidance is the how it can be done. The guidance information reinforces the concepts, describes various methodologies that can be used, and gives examples and references. It also gives information on the benefits and drawbacks of various methodologies. The guidance reports are not intended to promote a single methodology or to be an exhaustive treatment of the subject material. They provide information and references so that the user can more intelligently choose and implement the appropriate methodologies given the user’s application and capabilities. [ ]

Jan 5, 2003

By P. Bouéry

This account of a contrivance which has been found serviceable in practice may be of interest to engineers, and especially to those engaged in hydraulic mining. In that process, one feature which seems at first sight to possess little significance, is highly important in its effects—namely, the regular supply of water to the " giants." The chief condition of this regularity is an even flow in the pipe-line from the penstock to the nozzle.' This, of course, requires that the penstock be kept full. Any considerable surplus of water, overflowing at the penstock, is a simple waste, unless it be otherwise utilized in the sluices. At all events, it does not affect the delivery of the nozzles. But a deficiency at the penstock may admit air into the pipe; and this may produce hydraulic recoil, injuring the line by incessant vibrations, giving rise to leaks, and even to blow-outs, and (when a partial vacuum exists at the time of the fracture) the collapse of the pipe. An accident of this sort in our main pipe-line, occasioned partly by a settling of the ground and partly by the presence of air in the pipe, led to the employment of the device here described. The line referred to is made of 30-in. pipe, successively reduced to 18 and 15 in. in diameter, for the supply of two giants. By reason of the causes above mentioned, the 30-in. pipe burst more than 1,400 ft. from the penstock, and, despite the air-valve, the 30-in. pipe was collapsed to such a degree that one length of it, next to the penstock, was squeezed together and positively swallowed by the main pipe, through which it traveled for 750 ft., piercing through a short elbow on the way. This remarkable result was undoubtedly the effect of the vacuum suddenly produced when the pipe burst.

Jan 1, 1907

In order to overcome some of the deficiencies associated with conventional fine coal cleaning methods, a novel process known as microbubble column flotation (MICROCEL?) was developed at Virginia Tech. This article provides an overview of recent commercialization efforts and in-plant testing of this technology. The results of the in-plant test work indicate that the MICROCEL technology offers several advantages for the upgrading of difficult to clean coals. The test data have also been used to develop guidelines for the design of coal flotation columns and to validate existing scale-up procedures.

Jan 1, 1991

By Dale A. Sketchley

The objective of Quality Assurance and Quality Control programs on advanced exploration projects is to ensure that data from sampling, assaying, and recording of geological observations are of high integrity for the purpose of obtaining reliable mineral resource and reserve estimates. National Instrument 43-101 Standards of Disclosure for Mineral Projects states that all scientific and technical reports must be prepared and certified by a Qualified Person in accordance with professional and industry standards following Mineral Exploration Best Practices Guidelines and Definitions, and Guidelines of the CIM Standards on Mineral Resources and Reserves. Additional information is given in Companion Policy 43-101CP and Technical Report Form 43-101F1. Although the Qualified Person must discuss and describe the Sampling Method and Approach; Sample Preparation, Analyses, and Security; and Data Verification; there is little specific information in the guidelines on Quality Assurance and Quality Control programs. Examples from the Oyu Tolgoi Cu-Au Project are given for optimizing sampling protocols, following proper sample preparation procedures, safeguarding sample shipments through chain of custody, ensuring quality of assaying with Round Robin monitoring programs, verifying transcribed data, and validating data with check assays to ensure high integrity for resource and reserve estimation. Oyu Tolgoi, also known as Turquoise Hill, is in the South Gobi desert, Mongolia, about 550 km south of the capital city of Ulaanbaatar, and 875 km northwest of Beijing, China. The project covers a major copper-gold system, which consists of the South West, South, Central, and Far North deposits, comprising high sulphidation and porphyry mineralization with oxide, supergene, and hypogene ore types. Exploration is done by multiple diamond drill rigs, and samples are prepared on-site and shipped to laboratories in Mongolia, Canada, and Australia for further preparation and analysis.

May 1, 2003

By M. M. Gilkey

This report on water requirements for the New Mexico mineral industry includes data on industrial operations closely related to mineral production. It contains general information concerning public water supplies at five principal cities and a short discussion of water needs for agriculture. Background data are presented regarding the geography, hydrography, and water laws of New Mexico, as well as precipitation and evaporation characteristics of the State. Most of the information on source, quantity, distribution, cost, and treatment of water was obtained in interviews with company representatives. Water systems at 46 operations are illustrated by schematic waterflow diagrams. The study revealed that the New Mexico mineral industry in 1962 used about 16 billion gallons of new water and reused 152 billion gallons, a total usage of 168 billion gallons. Consumption amounted to 7.6 billion gallons. Of the 16 billion gallons of new water, 11.9 billion gallons was "self- supplied" from ground water sources, and 2.6 billion gallons was self supplied from surface sources. Approximately 1.5 billion gallons was purchased. Some of the water from company-owned wells is piped as far as 30 miles. Reported costs of the self-supplied new water range from 1 to 20 cents per 1,000 gallons and average 8 cents per 1,000 gallons for power and maintenance. Inadequacy of the supply of new water at many operations necessitates large-scale recirculation. Reported costs for recirculating water (power and maintenance) average 1.8 cents per 1,000 gallons. For the entire New Mexico mineral industry in 1962, the value of product was $42 per 1,000 gallons of new water intake and $88 per 1,000 gallons consumed. Analyses of 40 water samples, listed in the appendix, show the wide range in water quality at plants throughout the State. Projection of the 1962 water requirements of the mineral industry indicate that the total demand for new water will increase from the 16 billion gallons used in 1962 to 24 billion gallons in 1980, a 50 percent increase; and to 36 billion gallons in 2000, which is 125 percent more than the 1962 intake of new water.

Jan 1, 1965

By M. M. Gilkey

This report on water requirements for the New Mexico mineral industry includes data on industrial operations closely related to mineral production. It contains general information concerning public water supplies at five principal cities and a short discussion of water needs for agriculture. Background data are presented regarding the geography, hydrography, and water laws of New Mexico, as well as precipitatton and evaporation characteristics of the State. Most of the information on source, quantity, distribution, cost, and treatment of water was obtained in interviews with company representatives. Water systems at 46 operations are illustrated by schematic waterflow diagrams. The study revealed that the New Mexico mineral industry in 1962 used about 16 billion gallons of new water and reused 152 billion gallons, a total usage of 168 billion gallons. Consumption amounted to 7.6 billion gallons. Of the 16 billion gallons of new water, 11.9 billion gallons was "self- supplied" from ground water sources, and 2.6 billion gallons was self supplied from surface sources. Approximately 1.5 billion gallons was purchased. Some of the water from company-owned wells is piped as far as 30 miles. Reported costs of the self-supplied new water range from 1 to 20 cents per 1,000 gallons and average 8 cents per 1,000 gallons for power and maintenance. Inadequacy of the supply of new water at many operations necessitates large-scale recirculation. Reported costs for recirculating water (power and maintenance) average 1.8 cents per 1,000 gallons. For the entire New Mexico mineral industry in 1962, the value of product was $42 per 1,000 gallons of new water intake and $88 per 1,000 gallons consumed. Analyses of 40 water samples, listed in the appendix, show the wide range in water quality at plants throughout the State. Projection of the 1962 water requirements of the mineral industry indicate that the total demand for new water will increase from the 16 billion gallons used in 1962 to 24 billion gallons in 1980, a 50 percent increase; and to 36 billion gallons in 2000, which is 125 percent more than the 1962 intake of new water.

Jan 1, 1965

By Manuel Gomez

Coking time was investigated as a function of coal properties, carbonization parameters, and the interactions between these variables. The results indicate that coal properties and carbonization conditions interact to influence the coking time, and these interactions may be determined quantitatively. Prediction equations were developed and used to simulate the effects of bulk density, flue temperature, and oven width on coking times for coals having a wide range of properties. The data presented provide additional information on the carbonization process to permit the coke oven operator to maintain closer control of coke production and coke uniformity.

Jan 1, 1968

By J. Contreras

In the period 2003-2004, the situation of the copper concentrate market was unfavourable for smelters and refineries, with low TC/RC values and euro-dollar parity affecting European companies negatively. Within this scenario, Atlantic Copper decided to develop a plan to increase its competitiveness and to improve its results in these market circumstances. The plan affected all departments in the Company and was based mainly on cost reduction in energy and manpower. A general analysis of the Complex's overall efficiency was carried out and, based on this analysis, two major Projects emerged, "Primavera" and "KF". Each Project consisted of smaller sub-projects which aimed at improving the areas which could perform better according to our general analysis. The paper describes how this entire project developed and the final results obtained by the Company.

Jan 1, 2007

By N. Ogbonna

A mathematical model for bioleaching at the agglomerate scale is developed. The agglomerate is defined as a unit volume of a heap which contains a solid phase (a size distribution of ore particles), a liquid phase (stagnant and flowing leaching solution) and a gas phase (flowing air and air pockets). Leaching at the level of an ore particle is modelled using a diffusion-reaction equation, and oxidation by attached and detached microbes is treated separately. With the aid of the mathematical model, the effect of particle size distribution on the bioleaching of an ore containing a mixture of chalcocite and pyrite, and a culture of iron oxidizers, is investigated.

Jan 1, 2006

By Benjamin Talbot

The problem of segregation and cavities in steel ingots is a subject which has given and is still giving metallurgists, engineers, and operators matter for serious consideration. This question has come more into prominence lately in the United States, due largely to the desire of railroad engineers to secure a better rail than they have obtained ill the past, as the service they demand is increasing in severity; and no doubt manufacturers will be able to meet what is required of them. The rails here are as good as those made in other countries, but the conditions of service and the extremes of climate are more Revere and consequently more breakages occur. Various reasons have been advanced to explain why rails may not be as good in quality to-day as in the past. Some engineers consider that modern methods of manufacturing, designed chiefly to obtain large output, tend to reduce the standard of excellence of more deliberate methods. Others think that 4-ton ingots are worse than those of 2 tons. Again, it is stated that the 100-ton heat in one ladle is too large, and is a step in the wrong direction in casting. My experience upon the question of the size of the ingot is that in rolling rails of from 85-to 100-lb. sections, the range of ingot is only practically such that the difference in the size does not help in the question of segregation, cavities, or blow-holes. I call attention to two ingots from work done at Pencoyd, Pa., 16 years ago. These ingots were bottom-poured, cast at the same time from the same center-runner, the one being 20 by 24 in. and the other 13 by 16 in. in size. The steel was acid open-hearth, and contained C, 0.43; P, 0.062; S, 0.069 per cent. At the same place in each ingot where seg-

Jan 1, 1914

By W. D. Dietzman

In view of the developments and recent interest in recovering heavy oil, this report was prepared to present general information on more than 2,000 heavy oil reservoirs in more than 1,500 fields in 26 states. Information given on each heavy-oil deposit includes the following items: Field and formation, location, average or range of depth, average API gravity of oil, aver- age thickness of formation, areal extent of field, total productive wells drilled in field, discovery date of field, field status in 1963, cumulative production to Jan. 1, 1964, and all known types of secondary-recovery operations initiated (active and terminated) Fn the reservoirs. The geographical location of heavy oil accumulations in the United States is shown by a map, and a brief discussion of 14 different areas is given in the text. As of Jan. 1, 1964, 75 percent of the heavy oil reservoirs in the United States had produced more than 10 billion barrels of oil and are estimated to contain over 90 billion barrels of oil in place. It is also estimated that if the deposits which have had little or no production history, such as the one in Missouri, were included, the total remaining heavy oil in place would be in excess of 150 billion barrels. To find economical ways of recovering this vast reserve of heavy oil, industry is developing technology in several types of thermal-recovery methods. Some of the general considerations and more suitable limitations in selecting a prospect to study fur thermal-recovery techniques are: Lithology, depth, viscosity of reservoir fluids, gravity of oil, reservoir thickness, oil in place, and permeability and porosity.

Jan 1, 1965

By Robert W. Thomas

This paper describing the mining practices at the Ray mines at Ray, Arizona, is one of a series being prepared by the Bureau of Mines on mining practices, method and costs in the various mining districts of the United States. At the present time (1929) all mineral Production of the Ray district comes from the Ray Mines of the Nevada Consolidated Copper Co. The ore body is being worked from two shafts, designated as No. 1 and No. 2. An undercut block-caving method of mining is used, variations of which are employed to fit different conditions. The present method has been developed by making improvements as experience was gained or when conditions changed. Refinements in practice are continually being made. The Ray mine was the first in the Southwest to use a caving system of mining on a large scale.

Jan 1, 1929

By P. M. Gruzensky

In view of some of the outstanding properties of chromium, considerable effort has been made in recent years to prepare the high-purity metal in ductile form. The object of this investigation was to obtain such metal by electrolysis of chromic acid solutions and to determine the conditions that yield optimum results. The apparatus is described, as well as the parameters that received consideration. Ion-exchange resins were employed to remove cation impurities from the electrolyte. Tensile specimens prepared from the electrodeposits by arc melting, press forging, swaging, and drawing, were ductile at room temperature. INTRODUCTION Metallic chromium displays an exceptionally high resistance to corrosion, oxidation, and wear at elevated temperatures. Although these properties are utilized to some extent when chromium is alloyed with other metals, such as in stainless steels, high-chromium alloys and pure chromium are not employed as engineering materials because of a serious lack of ductility and impact resistance at room temperature. Because high-chromium alloys appear promising as refractory metals for applications such as in gas turbines, much interest has been shown recently in the preparation of uniformly ductile chromium and in determining which factors contribute to its brittleness.

Jan 1, 1957

By J. L. Wiley

[NOTE. - P before the number indicates that the ab tract appears in patent section. Abstract numbers without. P appear in literature section. A Abietic acid, hydrogenolysis - 404 Acetone, recovery - P-4024 Acetylacetonate catalyst - P-3352, P-3664, P-6067 Acetylene, production - P-2556, P-3169, P-4064, P-4221, P-4222, P-5002 Acid catalyst - P-4584 Salts - P-4555 See also Specific acid. Acid treatment. See Procedure. Adsorption. See Hydrogen, adsorption. Agitation. See Equipment, agitators; Procedure. Alcohols: Decomposition -1128 Hydrogenation - P-3288, P-4743 In motor fuel - 1486]

Jan 1, 1952

By Gerardo Muñoz

All along this century CODELCO's smelters have been operating on conventional technology, which Codelco has streamlined by using oxygen, developing oxy-fuel burners for reverberatory furnaces and introducing oxygen smelting of concentrates in converters (Teniente Converters). Chuquicamata began capturing sulfur from its first Teniente Converter in 1985. Since then, it has implemented a program allowing to gradually increase such capture. In 1994 a Frame Law on the Environment was enacted in Chile. By virtue thereof and of CODELCO's corporate environmental policy, an investment program aimed at capturing more than 95% of gaseous pollutants in all its smelters in a term not exceeding twelve years was established. The above entails a substantial improvement of the operational management as regards fugitive gases capture and treatment. This paper describes the way CODELCO is to deal with this challenge with a view to fully complying with environmental regulations and obtaining the required environmental certifications of its operations.

Jan 1, 1996

By Neil J. Sullivan, Goode. George E.

"The Adams mine, operated by Jones & Laughlin Mining Company, Ltd. and located near Kirkland Lake, Ontario, produces one million tons of iron ore pellets per year. This paper briefly describes the mining and concentrating operation and deals mainly with the operation of the pellet plant. The orebody consists of fine-grained magnetite and some hematite, with alternating layers of chert, and is mined by conventional open-pit methods. After crushing and grinding, the ore is concentrated by magnetic separation and flotation. Agglomeration is accomplished by the grate-kiln pelletizing process. The plant layout and general operation will be discussed in detail."

Jan 1, 1968

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By Zhi-yong Liu

A novel process, which includes hydrochloric acid leaching, iron powder cementation, sulphide precipitation and spray pyrolysis, was proposed to treat iron concentrate recovered from zinc kiln slag for comprehensive utilization. Treated by two-stage hydrochloric acid leaching, the leaching extents of Ag, Pb, Cu, Fe and Zn are 96.78%, 95.14%, 97.72%, 94.51% and 87.74%, respectively. More than 99% Cu and Ag are recovered from the leach liquor when iron powder is 25% higher than the stoichiometric requirement. More than 96% Pb and Zn is removed with three times of theoretical consumption of FeS, and the concentration of impurities of the final solution is less than 500 mg/L. Fe2O3 powder, which is spherical with a mean size of 12 μm and purity of 99%, is prepared with final solution at 700 °C. In this process, not only the metal values can be recovered effectively, but also the iron resources of zinc kiln slag can be converted into high purity Fe2O3, thus, the comprehensive utilization of iron concentrate recovered from zinc kiln slag can be realized.

By J. M. Tannenbaum

A depth-sensing micro-indentation technique is developed for evaluating mechanical properties of materials at elevated temperatures (to 1200°C). System description and design considerations of the test apparatus are described. Based on Hertzian spherical indentation approach, the micro-indentation test apparatus is capable of determining Young?s modulus and creep strength of test materials with flat, tubular, or curved architectures from room to 1200°C under either air or controlled gaseous environments. Calibration tests were conducted on a H13 Tool Steel coupon from 22°C to 500°C and single crystal Haynes 230 alloy at 1000°C with excellent agreement to the literature values. Indentation creep tests of Haynes 230 alloy were carried out at 1200°C with preliminary results showing good correlation to creep exponents. Further system refinements will be investigated, and additional testing will be conducted.

Jan 1, 2010