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By J. H. McNamara

A process has been developed for the successful hydrometallurgical extraction of copper from sulfide ores. The extracted copper in the form of a salt is converted by reduction to copper metal. Purification of the metal is described which includes a description of the casting of ETP quality copper wire bars.

Jan 1, 1978

By Harold Cook

THE purpose of this paper is to state briefly the inspection requirements of the Bureau of Ordnance, the specifications governing the inspection, and the physical and chemical properties of the steel used in the construction of ordnance for our Navy. It would be of interest to evoke discussion on points in these specifications which do not tend to insure the production of the very best material for the purpose intended. In other words, what requirements of our specifications are useless and also in what manner are our specifications deficient? It might be well at the, start to answer the first criticism of this paper which will probably be made, and that is that the title does not agree with the subject matter. Considering metallographic work as covering only the macroscopic and microscopic examination of metals, this paper will have gone beyond the subject; but it is believed that a broader definition of metallography is that it is the study of the relation of the internal structure of metals and alloys to their composition and to their physical and- other properties. The internal structure of steel is dependent on its physical and its chemical properties. These physical properties are dependent on the chemical composition and the mechanical and thermal treatment of the steel. So it seems that this subtitle would not be entirely necessary and that the title "Metallography of Steel for Ordnance Purposes" would cover a study, not only of what we can learn of macroscopic and microscopic examination, but broadly of the whole subject of the specifications, the fulfillment of the requirements of which produce the structural conditions revealed by macroscopic and microscopic examination. .

Jan 2, 1916

By G. A. Smyres

The Bureau of Mines investigated a HCI-02 leaching procedure to recover Cu, Ni, and Co from a low-grade bulk sulfide flotation concentrate prepared from the Duluth gabbro complex. Since the Duluth gabbro complex is the largest known and unexploited Ni resource in the United States, the development of options for extracting the metal values from the flotation concentrate of this domestic resource is necessary. Conditions were determined for producing a pregnant solution low in Fe to facilitate recovery of the metal values. Leaching experiments were conducted at 15 pct solids, 1000 to 1200 C, 0.2 to 0.3 g HCI per gram concentrate in a 50-gal reactor pressurized to 50 pSig with 02. Base metal extractions from the concentrate were, in percent, 90 to 92 Cu, 99 Ni, 97 to 99 Co, and 0.1 .Fe. Precious metals extractions were erratic and unpredictable; at best their extractions were, in percent, 55 Au, 90 Ag, 40 Pt, and 45 Pd. The reaction time at 1200 C was 0.5 h. The use of compressed air instead of 02 as the oxidant was also investigated. Solvent extraction using LIX 622 was used to remove Cu from the pregnant solution. Cementation on Cu was tested for precious metals recovery, while sulfide precipitation was used to produce a Ni-Co product. A conceptual flowsheet for treating the Duluth gabbro concentrate is proposed.

Jan 1, 1985

By D. F. Malan, J. A. L. Napier

This paper describes some important aspects associated with historical underground measurements in South African gold and coal mines. Deformation measurements were used to confirm the use of elastic theory to simulate the rock mass behaviour in the Witwatersrand gold mines in the 1960s. Although a prominent time-dependent component of stope closure was measured as early as the 1930s, it was ignored owing to the benefit of adopting elastic theory. Neglecting the time-dependent response of the rock for many decades resulted in important aspects such as the effect of mining rate, the effect of advance per blast, and the need for enhanced design criteria not being explored. Recent work is only now starting to address this gap in knowledge. In-situ measurements of large coal specimens in the 1960s and 1970s indicated that a linear formula may possibly be a better approximation of coal pillar strengths. This alternative formulation was never adopted, however, as the power law strength formula was already deeply entrenched in the industry at that stage. In spite of these apparent failures to continuously generate and adopt new knowledge, a key lesson learnt is that major advances in rock mechanics will not be possible without careful monitoring of the rock mass behaviour in experimental sites. Areas requiring further research, such as pillar strength formulae for the Bushveld Complex and enhanced design criteria for the gold mines, can only be developed using extensive undergroundmonitoring programmes.

May 1, 2021

By D. Symonds, A. J. Richins, M. G. Nelson

Mining activities at Kennecott and Park City are similar in that both districts began operation in the last half of the 18th century, and both operated on a large scale for many years. However, they are different for two important reasons. First, the mines in the Park City district were always underground. The city of Park City was above many of those mine workings. It remained in place after mining activity stopped, and as the town became an up-scale ski resort, property values increased and the tailings dumps became desirable building sites. In contrast, Kennecott eventually became an open pit mine, and the growing pit effectively consumed many of the small towns that had grown up around the underground workings. Even towns on the periphery of the open pit were razed to allow for the disposal of mine overburden. Second, almost all mining activity stopped in Park City in the early 1970s, just as the first rigorous environmental legislation in the United States was enacted. (The Ontario mine operated sporadically until 1982.) Many mine structures, waste dumps, tailings ponds and piles, and tunnels were simply abandoned as the mining companies ceased business or were bought out. Kennecott, on the other hand, has been in continuous operation for more than 120 years, so a closure plan hasn’t been finalized. Nonetheless, environmental and social problems have arisen, and the company has had to address them.

Mar 2, 2022

By Andrew Chapman

The scope of the SA (Situational Awareness) working group is to analyze the effectiveness of information delivery systems in mining equipment and develop a unified display design, to mitigate against the proliferation of screens within equipment cabs. This working group comprises a variety of stakeholders representing a cross-section of relevant mining and mining-related organizations. This stakeholder network will help guide the direction of the working group towards a common, mutually beneficial end. For now, the focus is on shovel displays; however it should be noted that this work is expected to form the foundation for similar work with other equipment in the future.

Aug 1, 2013

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By W. J. Murray

Tests were conducted during the middle of this year to evaluate techniques and the equipment used in a novel continuous mining system for use in narrow reef tabular orebodies. Equipment was designed and supplied to test a mechanized drilling rig, a selection of rock breaking techniques and an integrated rock removal system, mounted on a self-advancing reciprocating flight conveyer. As the rock breaking techniques do not depend on the use of large quantities of conventional explosive the systems and equipment employed here have the potential to run continuously for 24hours per day. Whilst the tests carried out use simple prototype equipment employing established technologies, the long-term objective remains the use of sophisticated machinery capable of running continuously and un-manned, only interrupted by brief service intervals. Keywords: Hydro-fracturing, Rock breaking, Low energy, Propellant, Non-explosive, Safety, Continuous mining, Rocksplitter, Boulder buster, Ro-bust.

Sep 1, 2001

By William Mayrsohn

IMTRODUCTION The marketing aspects of a new precious metals property is often not considered until the mine actually begins production. At this point in time it is often too late or too expensive to change the mine's final product and not enough time to do an effective marketing program. MINE PRODUCT The rule for determining the mine's final product: The higher the gold and silver content, the greater return of metal at the lowest post-mining cost. . Dore Bullion: Most advantageous Advantages: - Highest return of metal. - Lowest refining cost per ounce of gold and silver. - Very compact and easy to ship in small quantities. - Easy to assay. - Assays between mine and refiner are comparable. - Melt losses are very small at the refiner. - Outturn of gold and silver 20-30 days for quick cash flow. - Gold/silver refiners can return or purchase metal soon after arrival at a nominal financing cost.

Jan 1, 1987

By Y. Zhen, N. Topliffe, G. Shou

The typical service life for a long distance slurry pipeline is15 to 25 years. Determination of wear rate for a pipeline could have a high economic impact on the pipeline project. Wear is a combination of chemical corrosion and mechanical abrasion. The contribution of each factor has to be determined project by project. Normally, chemical corrosion dominates the pipe wear in fine slurry pipelines. These pipelines have been designed with a velocity between 1.5 m/s and 2.5 m/s. Solids are suspended with high homogeneity across the pipeline section. Bottom wear is not much greater than the top wear and not sensitive to velocity. The wear pattern in the pipeline transporting medium coarse solids, such as milled ore and tailings, is different. Wear concentrates on the bottom of pipeline and the outer bends, and is more sensitive to flow velocity. Wear distribution along the pipelines is similar.

Jan 1, 2007

By C. J. Newman

Prior to 1995, the smelting process at the Kennecott Utah Copper Corporation (KUCC) smelter consisted of Noranda reactors and Pierce-Smith converters. In the consideration of expansion opportunities, the existing process did not lend itself to a program geared towards processing the total output of KUCC mining and milling operations. During 1989 and 1990, a number of smelting options were evaluated. The final selection was the Outokumpu flash smelting process, coupled with the new Kennecott-Outokumpu flash converting technology. Facilities were designed and installed to process 1:0 million tonnes per year of copper concentrate, with the ability to expand to 1.2 million tonnes per year. Following the shutdown of the smelter facility utilizing the Noranda technology in May 1995, the flash smelting, furnace commenced operation in June with the flash converting furnace startup following a few weeks later. The smelter has achieved design production targets and environmental goals. Current operating rates are at or above design rates for concentrate and matte processing. Emissions are lower than the permit levels and equate to less than 3.5 kg of SO2 per tonne of copper. This paper outlines the design, construction, startup and the first two years of operation of the modernized KUCC smelter.

Jan 1, 1998

By T G. Callcott

Australian coking coals are either class 434<tlj534<2l, at Port Kembla, or 634<3> at Newcastle. The former yield good blast-furnace cokes; the latter yield poor ones. However, increases in conventional strength indices, by methods applicable with normal coking ovens, seem unlikely to be economically attractive. Characterization of the coals has enabled an "economic" blend to be developed in the Newcastle area, and this has affected mining, preparation and handling operations. Coking studies have led to a specification for the making of strong coke, especially from the high-volatile Newcastle coals. Continuous coking studies, developed from the efforts to make very strong coke, have taken two directions. Firstly, coking thin layers of coal on strands; secondly, coking, in various types of vessels, compacts of coals with or with-out the inclusion of non-coking coals, iron ores or other inert materials. Gas, "sand" and hot metal coking have been tested for use in compact coking. Particular attention has been paid to pellets of coal and ore. The probable industrial problems in adopting any of the ideas are presently being examined.

Jan 1, 1963

By J. E. Rourke, O&apos

This paper describes the use of index testing to obtain rock properties that are useful in the design and construction planning of an underground development for civil engineering or mining projects. The index properties discussed include: point load; Schmidt hammer hardness; abrasion hardness; and total hardness. The first two index properties correlate to uniaxial compressive strength (UCS) and Young&apos;s modulus. Discussions are given on empirical, normalized relationships of UCS to rock mass properties and the integrated use with semi-empirical, geotechnical design methods. The hardness property indices correlate to construction performance parameters and some relevant experience is cited. Examples of data are presented from an index testing program carried out primarily on siltstone, sandstone and limestone rock core samples retrieved from depths up to 1005 m (3300 ft) in a borehole drilled in the Paradox Basin of eastern Utah. The borehole coring was done for a nuclear waste repository site investigation.

Jan 1, 1990

By Boyes B. A, Haydon D. J

The Paddington project is a gold mining and processing operation located 35 km north of Kalgoorlie in the Eastern Goldfields region of Western Australia (Figure 1). The project, which is operated by Pancontinental Goldmining Areas Pty Ltd, treats 1.25 Mtpa of oxidised gold ore, derived mainly from open cuts adjacent to the plant, with minor tonnages from other pits in the region treated on a custom-milling basis. Gold is recovered using a conventional CIL process. The process raw water supply of 3 500 kUd is derived from the local groundwater system. Average salinity of the raw water is around 90 000 - 100 000 mg/L total dissolved solids. A small portion of the raw water is passed through a vacuum distillation desalination plant to produce the project&apos;s 200 kUd requirement for high quality water for potable use and gold recovery.

Jan 1, 1988

By J. V. Richards

(Canal Zone Meeting, November, 1910,) THE Altar district, State of Sonora, Mexico, is for the most part a desert with but little rain-fall and few running streams. On account of this scarcity of water it is necessary for the natives to " dry-wash" the placer-gravel, and the object of the present paper is to describe some of the more common Mexican dry-washing devices. My observations were confined to a property known as Las Palomas placers, 15 miles from the Gulf of California and about 150 miles south of the United States border; a brief description of which property may be of interest as illustrating what is, perhaps, a typical Sonora dry placer. It is said that as long as a century ago the Yaqui Indians discovered rich cement-gravel, known as argo masa, at Las Palomas, and that there has been a more or less steady production of gold from the property ever since. The argo masa is a fairly-coarse gravel, the pebbles being quite firmly held together with a lime cement. Gold is scattered through this cement in particles ranging from fine specks to pieces several grams in weight, and occasional large nuggets are encountered. The contents vary so widely that it is difficult to estimate the average value, but returns of from $10 to $50 per cubic yard are common, and it is doubtful if the Indians considered anything less than from $4 to $5 a yard as "pay dirt." This argo masa occurs in channels varying in thickness from a few inches to several feet, and in width up to 300 ft.. These channels are continuous and are probably ancient stream-beds. They are for the most part covered with an over-burden of a more recent gravel, varying from 30 to 50 ft. thick, and the fact that this over-burden contains values as high as $1.50 per cubic yard, and yet was not worked by the Indians, shows that their idea of what constituted pay-ground was high.

Apr 1, 1911

By Hayward P. G

Complex ores, characterised by intimate intergrowths between minerals, have previously been treated by flotation with limited success. The most common problem has been that the liberation size is often finer than the lower size limit for selection flotation. A technique has been developed in which the problem is overcome by the use of a selective agglomeration procedure in which the values are coalesced with oil to a size suitable for flotation prior to introduction into the flotation machine. A complex lead ore was characterised by QEM*SEM mineralogy as having a mean galena liberation size of 15 ¦m. The best recovery of lead at 50 per cent Pb grade by standard flotation was 53 per cent. After application of the procedure, 86 per cent recovery was obtained after a single cleaner step. The oil functions as a bridging liquid in the presence of conventional collectors which render the mineral oleophillic.

Jan 1, 1991

By B. Seegmiller, W. A. Hustrulid, O. Stephansson

Introduction The concept of in-the-wall haulage was formally introduced to the mining community in the early 1970s. The basic concept is simple. Rather than the traditional in-the-pit haulage/access ramp, single or twin ramps driven in the pit wall are used with crosscut connections on the working levels. The in-the-wall pit haulage system could be considered as part of the opening of a new mine, a way of completing to final depth an existing pit, or part of the transition from open-pit to underground mining with the ramp remaining as part of the access. This paper is intended to acquaint the reader, particularly mine planning engineers, with some of the possibilities presented by the concept. Nordstrom (1983) said, "In many countries, mining engineers are either open-pit or underground specialists and, hence, often don&apos;t consider the possibilities for combining the best aspects of each into a hybrid mining system." This is believed to be true. There are few "hybrid mines" employing, for example, open-pit methods with underground haulage systems. One reason is the variance with the experience of the mining company and the engineers involved. However, with today&apos;s low metal prices, innovative and unconventional thinking may be just what a property requires to remain in production. All the possible situations where an in-the-wall ramp might be considered cannot be presented here. Rather, a few examples will be discussed for illustration purposes. Potential savings in stripping costs A major cost in open-pit mining is associated with the removal of overburden and waste rock. Although much of this stripping cannot be avoided, the volume (V) associated with the presence of an in-pit haul road can be quite large, V 100 (pit depth) a (road width) 2 x road grade (%) As an example, consider a vertical ore body 400 m (1312 ft) long, 60 m (197 ft) thick, extending to a depth of 160 m (525 ft). Cross sections are shown in Fig. 1. Other data are: ore tonnage = 11 Mt (12 million st), ore density = 2.9 t/m3 (0.09 st per cu ft), waste rock density = 2.7 t/m3 (0.08 st per cu ft), road width = 20 m (66 ft), road gradient = 10%, and level interval = 10 m (33 ft). The evaluation has been made assuming that the pit slope (0) is constant around the pit except on the side with the haul road. On this side, the slope is decreased to allow for the road. Slope angles of 30° to 60° have been selected to indicate the dependence of stripping volume (V) on slope angle. The stripping at the ends have been calculated assuming representation by cones. The volume associated with the haul road is 2.6 Mm3 (92.8 million cu ft). A comparison of the strip volumes involved is shown in Table 1. As can be seen, the percent savings in strip volume varies from about 4% to 17% and depends greatly on the slope angle. It is perhaps more interesting to examine the savings in stripping costs that could then be used to cover the cost of the access ramp and the crosscuts for stripping costs of $3/m3 ($0.08 per cu ft). The cost with the road volume is 2.6 million x 3 = $7.8 million.

Jan 1, 1988

For more than 15 years after the Northparkes orebodies were first discovered in 1976, the project was delayed by the need to find an economic way of mining the low-grade E26 deposit. The decision to use a low-cost block caving method to mine the orebody was made in the early-1990s; but the project&apos;s future did not solidify until 1993 when a radical new approach was adopted which sought to integrate business imperatives with technology and organisation strategies. The approach led to the establishment of a strategic vision which tied together all the components that could have a major impact on the project, to provide a framework for guiding strategic choices. The strategic vision strived to break away from the old rules of how mines are organised and operated; and involved finding imaginative new ways to radically redesign mining processes to achieve quantum leaps in performance.

Jan 1, 1997

By Yoginder P. Chugh

The control of soils has been found to be one of the highest priority factors influencing the return of reclaimed prime farmland to production. The compaction of soil is primarily a result of the movement of high ground pressure wheel scrapers and dozers over soils during the soils replacement and regarding process and occurs to a maximum depth of about 1.2 m (4 ft) The deleterious effects of soil compaction can be controlled by deep soil loosening to a depth of 1.2 m (4 ft). Most farm tillage equipment can only loosen soils to a maximum depth of 0.33 m (1.0 ft). Deeper soil loosening equipment, in the 0.33-1.2 m (1.0-4.0 ft) range is currently avail- able in the U.S. and abroad which may have potential for application on reclaimed prime farmland. This study has appraised such available equipment and the results indicate that a few manufacturers in the U.S., West Germany, and England market static and vibratory soil loosening equipment which have potential for deeper soil loosening of reclaimed prime farmland.

Jan 1, 1985

By John I. Yellott

SINCE the days of Newcomen and Watt, when men first sought to turn the energy of fuels to useful purposes, coal-generated steam has supplied most of the power needed for both stationary and mobile applications. The low efficiency and large water requirements of the nineteenth century steam-power plant caused many engineers to turn their thoughts toward more direct means of converting heat into power. As far back as 1791, John Barber secured an English patent on a combination&apos; which he might well have called a "gas turbine." He proposed to generate procducer gas from coal or charcoal, to burn it under pressure in air obtained from a reciprocating compressor, and to drive a turbine wheel with the resultant hot combustion products. With the optimism characteristic of the inventor, he hoped that the turbine would not only operate his compressor but also perform useful work. There is no record that Barber ever built his machine, but if he had done so he would have learned that useful power can result from such a combination only when both turbine and compressor are highly efficient and when the temperature of the gas entering the turbine is not far below 1000°F. Hot-air prime movers continued to attract attention, and both Stirling in England and Ericsson in America2 built reciprocating engines that operated with some success upon coal-heated air. The internal-combustion engine, burning liquid fuels, supplanted the hot-air type, and, for the first time, challenged the supremacy of both the steam cycle and its familiar fuel, coal. LIQUID FUELS COMPETE WITH COAL As the increasing use of the automobile multiplied the demand for gasoline, the supply of residual fuel increased rapidly, and heavy oil declined in price to the point where it became competitive with coal. First in the marine field and later in stationary plants in areas adjacent to refineries, oil began to be widely used. Oil&apos;s greater convenience and higher heating value per pound of weight or per cubic foot of bunker space have caused it to take over much of the marine field. The introduction of the highly efficient diesel engine, capable of burning a fuel that is cheaper than gasoline, again threatened the position of coal, for the operating costs of the diesel became directly competitive with those of coal- burning equipment. In isolated power plants and smaller marine installations, the diesel provided the medium by which oil again successfully invaded fields formerly dominated by coal. The diesel engine, provided with an

Jan 1, 1946