Search Documents

By C. L. Elmore

This paper describes a new development in the cyanidation of gold ores, which we expect will result in a major improvement in the way many gold ores are treated. The process has evolved through work at Hazen Research, Inc. for Kamyr, Inc. and in brief, the consists of utilizing oxygen rather than air in the carbon-in-1 each (CIL) process. We refer to this improved process for all ores amenable to cyanide leaching as carbon-in-1each with oxygen or CILO.

Jan 1, 1988

By B. S. Crocker

Introduction Pebble mill grinding can be defined as wet or dry grinding in a tumbling mill with pebbles used as grinding media instead of metal balls. To clarify the terminology, which may differ in the various mineral industries, it is essential to point out the differences between grinding with partially sized and carefully sized grinding media. When grinding with partially sized media (limited top size only), sizing is usually the setting of the primary crusher and there is no bottom sizing-all fines, chips, sand, and small sizes are included. This is often called run-of-mine or primary autogenous grinding. In this case the entire crusher discharge is fed into the mill and part of it becomes "charge" and the rest is "sand feed." When grinding with carefully sized media (i.e., media that is sized both at the top and bottom of the size range), no fines or chips should be present in the charge to the mill. This is pebble milling and the grinding media consists of ore, rock, Danish pebbles, gravel pebbles, or ceramic pebbles-media that is appreciably lighter than steel balls. The grind¬ing charge is fed to the mill to maintain a load and it does the grinding usually on a classified sand product of suitable size. If the grinding media is the same as the raw material being ground, this type of grinding is also called secondary autogenous grinding. Pebble milling was in common use in the gold mines of South Africa in the early 1900s. It is by far the oldest form of autogenous grinding. It may be used to grind crushed ores from a maximum size of about -% in. to all -10 µm. Pebble milling differs from "run of mine" grinding in many ways, but two big differences are: I) The ore itself is crushed to approximately t/4 in., or it may be further reduced to 8 or 10 mesh in a rod mill or fine crusher. 2) Then carefully screened ore of the correct size is placed in one or more stages of pebble mills to grind the ore to its desired final size. The media size is carefully selected to be the most efficient for the particular size reduction required in each stage. This is a precise practice, and pebble mills can be designed as easily and accu¬rately as can rod mills or ball mills by those skilled in the art. A pebble mill is a rotating cylinder in which the grinding media has been carefully sized for the job in hand. If the mill is doing coarse grinding, the media will be correspondingly large and the mill might be called a primary (or intermediate) pebble mill or a lump mill. If it is to do fine grinding, the media selected is proportionately smaller and the mill would be called a secondary or tertiary pebble mill. The media itself is normally rounded pebbles made from screened ore from the primary-crusher discharge. The charge could be rounded pebbles of the correct size removed from a primary autogenous mill, or the pebbles could be carefully sized gravel, sized beach pebbles, or sized flint pebbles. Usually the specific gravity of a pebble is from 35 to 55% of that of steel balls. It commonly varies from 2.66 to 4.2, depending on the specific gravity of the ore. Primary pebble mills, or lump mills, would be charged with pieces of rock up to 10 in. maximum and 5 in. minimum and would grind crusher discharge with a maximum size of 1/4 in. on hard ores and 3/8 in. on softer ores. Secondary pebble mills could be charged with pieces of rock screened on 4 x 4-in. square mesh and retained on 1 1/2 x 1 1/2 in. These mills would grind material with a top size of 8 to 10 mesh. Tertiary pebble mills could be charged with screened rock that would pass 3 x 3 in. and be retained on 1 x 1 in. and grind 35 to 48 mesh sands. Principles of Operation Media Size To operate a pebble mill efficiently it is imperative that great care be taken in screening the rock feed that is fed into the mill to round up into a range of pebble sizes that are suitable for grinding of the sands that are also being fed into the mill, having regard to the top size and the average size of such sands. Since most mill operators know from experience the correct size of steel balls to be used in a mill to handle different sizes of sand feed, it is convenient to relate the pebble size to a ball size. If the correct ball size is not known, it is relatively simple to conduct batch laboratory tests in 12- to 18-in. laboratory mills to determine the optimum size of steel ball. Also, Bond and others have published formulas for calculating the optimum ball size. When the correct size of the individual steel ball is known, then a pebble of the same weight will perform the same grinding if used in the proper size of mill and under the proper conditions, which will be defined later. This media size, as defined by the average weight of the individual pieces, is most important. The following tests will serve to clarify this point. Laboratory Tests Using Steel Balls or Pebbles. Several labora¬tory grinds were made to compare a given size and weight of steel ball with a pebble of the same weight but larger size, and of the same size but lighter weight. The following loads were used in the 12-in. laboratory mill: All 1 1/4-in. steel balls weighing 122 ± 10 g. All 1 3/4-in. pebbles weighing 123 ± 5 g. All 1 1/4-in. pebbles weighing 45 ± 3 g. To equalize capacity, the test with the steel balls was run only 34% as long as the tests with pebbles. In each case, 2,500 g of primary sands were ground at the optimum gravity for the media being tested at a speed 78 1/2 % of critical speed, and for 170, 340, and 680 revolu¬tions for the steel balls, and 500, 1,000, and 2,000 revolutions for the pebbles. The first test is summarized in Table 74 by showing the 80% passing size in microns for the feed and for each of the grinds. Two other tests are summarized in Table 75 showing the screen analyses

Jan 1, 1985

By W. C. Dailey

The Groveland Mine. Concentrator and Pellet Plant are located in Dickinson County in the Upper Peninsula of Michigan. The construction of the Groveland Mine in 1959 was the first United States operation for the Hanna Mining Company in mining and beneficiating low grade iron formations. (1) The plant produced concentrates at 675,000 lt/yr. (746,600 mt/yr.) which was used as sinter feed in the steel industry from 1959 to 1963. Upon completion of a pelletizing plant in April, 1963 production increased to 1.6 million lt/yr (1.77 million mt/yr) of pellets. After the plant expansion, mill capacity increased to 2.2 million 1t/yr (2.43 million mt/yr.), producing 2.0 million It/yr (2.23. million mt/yr.) of pellets assaying 63.0% iron and 7.0% silica. (1) Prior to April, 1971 the concentrator operated with a flowsheet consisting of spirals and a petroleum sultonate iron notation circuit. This was presented in a paper by R. R. Smith and E. E. Sougstad, "Flotation at the Groveland Mine" in the 50th Anniversary Volume entitled "Froth Flotation". The ore body at the Groveland Mine consists basically of two types of iron formation (34% total iron, 20% magnetic iron head), the most abundant of which is a bedded magnetite-chert-silicate rock. This is intermingled to varying degrees with non-bedded to poorly bedded hematite-clastic quartz-chert rock. (1) In developing the Groveland flowsheet, Humphrey spirals were used because of their ability to produce a coarse size concentrate at a low cost. Tests run in the lab and pilot plant on the Groveland ore showed a. spiral weight recovery in the 27 to 33% range. (2) However as the ore body was opened up, it became evident that the most abundant ore type did not liberate at the coarse size fraction needed for spirals. Plant recoveries for the first five years operating the spirals only averaged from 10 to 15% weight recovery. In 1971 the weight recovery on the spirals had dropped to a low of 5%. To increase iron recovery and at the same time increase plant capacity with the least effect on the present operations the mill flaw-sheet proceeded from a spiral-notation circuit to an all flotation circuit to a magnetic separator-flotation circuit. During this changeover the concentrate grade also had to be improved to meet market demands. To do this the following changes evolved

Jan 1, 1978

By None None

Use this month wisely — go online and take advantage of your new member benefit. As an SME member, you can read and download for free all of the full-text papers of the Mining, Metallurgy & Exploration (MME) journal, including the 22 papers in the launch Special Issue on Mineral and Metallurgical Processing in Honor of Professor Emeritus Douglas W. Fuerstenau as well as the archived articles of Minerals & Metallurgical Processing (MMP), all the way back to Vol. 1, No. 1. Just follow the instructions on this page.

By Terry S. Maley

Federally owned mineral deposits generally fall into one of three categories: (1) minerals locatable under the General Mining Law of 1872; (2) minerals leasable under the Mineral Leasing kt of 1920; and (3) minerals salable under the Materials Act of 1947. Unlike valuable deposits of gold, silver, copper, iron, lead, etc., which may be locatable, industrial minerals are acquired through a variety of methods depending on use, composition, and date of acquisition. Table 1 may be used to establish the method of acquisition for most industrial minerals.

Jan 1, 1987

By Mikko Ruonala, Pertti Lamberg, Olli Hyvärinen

The smelting of copper sulphide concentrate is the dominant route in copper production (Figure 1). Heap leaching of oxidized copper ores with weak sulphuric acid together with copper solvent extraction and electrowinning has gained popularity during recent decades, and it accounts for close to 20 % of primary copper production. Heap leaching offers advantages such as lower costs and consequently the construction of a production plant at a mine site is made possible. Limited resources of oxidized copper ores, the long leaching time of chalcopyrite and inability to recover gold inhibit this growth. The HydroCopperTM process developed by Outokumpu Technology at the laboratories of Outokumpu Research, Pori, Finland, is a new chloride leaching technology /1,2/, which offers an alternative to smelting and heap leaching (Figure 1). In the HydroCopperTM process sulphidic concentrates with various mineralogy and grade can be treated at the mine site to produce high-quality copper. Instead of copper cathodes the HydroCopperTM end products are copper wire rod, copper bars or copper billets. One of the advantages of HydroCopperTM technology is that gold can also be recovered in the process simultaneously /3/. To confirm the functioning of the new process and to dimension equipment Outokumpu Technology built the HydroCopperTM demonstration plant in Pori, Finland /4/. It has a capacity of 1 metric ton of copper per day. After testing with Pyhäsalmi copper concentrate in 2003, the first commercial demonstration run with Erdenet copper concentrate took place in 2004.

Jan 1, 2005

By Mohamed Gamal, Bill Zietlow

"In oil fields that have been depleted with conventional surface oil wells but still contain significant reserves, an extraction alternative that has been constructed recently consists of a drilling gallery below the oil-bearing formation at the bottom of a 1,200-ft deep shaft where upward sloping wells are being drilled to allow oil collection by gravity. The gallery is 92-ft diameter with a flat ceiling supported by rock bolts and shotcrete structurally connected to the shaft liner, which doubles as a large tension support element. A special excavation sequence was implemented for the gallery to maximize arching in the stress above the large diameter gallery. The gallery was located in competent sandstone, constructed with sequential excavation methods, and designed with a 3D finite element model. Special consideration was given to optimizing the gallery location to take advantage of the geology for construction and support while giving good positioning to the well drilling equipment. BACKGROUND Brierley Associates was contracted to provide both design and construction management services for an approximately 1,165-ft deep, 18-ft excavated diameter vertical shaft that opens to a 92-ft diameter flat roof gallery. The geotechnical data for the site was based on information available from literature and from data obtained during core drilling and from in-situ testing in a 1250-ft deep borehole. Wire line coring was conducted using a HQ, 10-ft long core barrel with a mud-rotary drilling method through sedimentary bedrock. The geometry of the shaft and the bottom gallery is shown in Figure 1(a). GEOLOGIC SETTING AND STRUCTURAL GEOLOGY The stratigraphy at the site includes a series of interbedded marine limestone, shale, claystone, siltstone, sandstone, and chert units. Limestone and shale are the predominant lithologies in the upper approximately 675 ft, while the lower portion contains five distinct sandstone units and more abundant shale, claystone, siltstone, and coal than the upper zone. A cherty sandstone at a depth of 1,130 below ground surface (bgs) was encountered. At a depth of about 1,000 ft to 1,116 ft bgs, a non-marine sandstone unit was encountered. This sandstone is the hydrocarbon bearing layer. Joints and fractures observed in the limestone had dips ranging from 0 to 90 deg. from horizontal with joint and fracture surfaces that ranged from planar to irregular and smooth to rough. Several, near vertical shear planes with slickensides were observed with surfaces that were generally smooth and planar. RQD values for the limestone ranged from 0 to 100 percent but in general were good to excellent."

Jan 1, 2016

By B. Pina

The conversion of maf coal to pyridine solubles and gas was investigated in a microautoclave batch reactor for two Kentucky washed coal samples (Ky. #9 and Ky. #11). Solvent hydrogen donor strength was controlled by the tetralin content of a mixed 1-methyl naphthalene-tetralin solvent. Primary liquefaction conversions greater than 90% in 10 minutes of reaction time at 427°C were achieved in liquefaction solvents with moderate hydrogen-donor strength (circa 40% tetralin). A modified irreversible first order rate expression, containing a parameter (a) to account for the intrinsic reactives of the maf coal, described the experimental kinetic data rather well. The integrated form of this rate expression is -k1e x = a (a- ß)e (eq. 1) Parameter ß in equation (I) is the pyridine solubility of the starting maf feed coal, x is the primary liquefaction fractional number of conversion, k1 is the specific reaction velocity constant and ? is reaction time. Primary liquefaction rate was found to be quite temperature sensitive and independent of coal when a liquefaction solvent containing 40% or more tetralin was employed. Primary liquefaction rate appears to depend on liquefaction solvent hydrogen donor strength for solvents containing = 40 wt.% tetralin. The intrinsic reactives in the maf feed coal or a appear to depend mainly on the inertinite group maceral content of the coal for the two coals studied in liquefaction solvents containing 3 40wt.% tetralin. In contrast, a appears to decrease in liquefaction solvents of low donor strength., i.e; solvents with less than 40wt.% tetralin.

Jan 1, 1984

By M. G. Mudrey

Considerable potential for metallic mineral resources in the lower Proterozoic rock of the Lake Superior region exists (Mudrey and Kalliokoski, in press). Most of the identified resources occur within a 2.0- to 1.9-b.y.-old lithotectonic assemblage characterized by extensive sedimentary strata in the north and thick volcanic se- quences in the south, all deposited on a two-terrane, composite basement of Archean age (DeMattis, 1989). Since the late 1960s, measured resources of nearly 90 Mt (100 million st) of typical volcanogenic massive sulfide at eight localities in Wisconsin have been discovered. The largest of these, near Crandon, contains about 61 Mt (67 million st) of ore averaging 5.6% zinc, 1.1 % copper, and 0.5% lead. Although none of these properties are in production, estimates by Mudrey (1978) suggest that Wisconsin could become a major contributor of newly mined zinc and copper in the United States. These discoveries and uranium and gold discoveries in the Canadian Precambrian are what have led to in- creased metallic mineral exploration throughout the northern half of Wisconsin (Fig. 1). In this chapter, we discuss the geology and legal framework in which the Proterozoic volcanogenic massive sulfide deposits were found. Exploration activities for other metals have generally been negative, and have been background noise for the legislative activities related to the massive sulfide discoveries. Uranium exploration has been consistently negative, although several different environments have been assessed. Other metal discoveries. such as titanium/vanadium, have been subeconomic.

Jan 1, 1991

By A. L. Mular

Optimization has been defined as the procedure of locating the best of something or the best way of doing something (Mular, 1971). Several texts (Himmelblau, 1972; Ray and Szekeley, 1973; Box and Draper, 1969; and Lee, Adams, and Gaines, 1968) on the subject are available. Such techniques have been placed in two categories, namely, off-line optimization and on-line optimization. For the former, an optimization model is employed; for the latter, a black box approach is involved. The essence of an optimization problem becomes obvious, when a response surface is considered. Suppose that a response (in this case, a criterion of improvement such as profit per day; grade of concentrate per day) denoted by Y, depends only on two variables, X1 and X2, which can be manipulated and controlled. The relationship between Y, X1 and X2 may be represented graphically as shown in Fig. 1. The diagram is called a response surface and its height, Y, is determined by X1 and X2. Points associated with equal responses constitute response contours. The shape is that of a hill, and optimization is the procedure of ascending the hill to reach the top (maximum). In practice, a response surface is subject to disturbances which distort its shape and shift its location from time to time, so that the exact location of a maximum is imprecise. Generally, constraints (restrictions of some sort) on responses and manipulable variables serve to define a region of permissible operation.

Jan 1, 1979

By H. Soto

As a follow-up to a laboratory study, a pilot scale column for the flotation of coarse particles was designed and then tested in two different applications. In the first test, the concentration of a coarse phosphate ore (normally beneficiated by skin flotation in spirals) was carried out in a flotation plant in Florida. In the second pilot test, the column was used as a scavenger cell in the coarse flotation section of a potash beneficiation plant in Saskatchewan, Canada. Particle size of the phosphate feed was of the order of 0.6-1.0 mm and the potash feed 1-3 mm. The column was 1 feet diameter and was fed at rates between 0.5 to 1.4 tph. The effect of operating variables such as air and water flow rates, bubble size, frother concentration, residence time, sparger, etc were evaluated. The most important factors in column performance were found to be bubble size, air flow rate and the thickness of the froth bed. Best results were obtained in the absence of a froth bed, with low air flow rates and with small bubbles. The paper presents comparative performance and discusses details of the column design. Phosphate recoveries in the column were better than 90% whereas the conventional skin flotation circuit normally yields about 60% recovery for a comparable concentrate grade. Potash recovery from the tailings of the conventional flotation circuit was as high as 42% recovery. Particle size distribution of the potash concentrate was coarse enough to meet requirements for potash fertilizer.

Jan 1, 1993

By J. L. Arola

Since start-up of fluxed pellet production in August: 1987, the Minorca Mine has made a number of process changes to improve pellet quality and productivity. Quality improvements include: 1. Installation of a contraction t:ester. 2. Installation of on-line calcium and silica analyzers. 3. Process autocation changes. Productivity improvements include: 1. A second fluxstone storage bin and feeder. 2. A new vacuum pump and disc filter. 3. A filter aid addition system. 4. Enclosing the filter cake reclaim feeder. 5. A new grate bar cleaner.

Jan 1, 1989

By G. V. Sullivan, B. E. Davis, T. O. Liewellyn

A technique was devised by the US Bureau of Mines to recover coarse [+1.2-mm (14- mesh)] North Carolina phosphate now discarded as waste. Concentration utilizes the difference in particle sphericity between the relatively flat calcite and the rounded phosphate pebbles. Sizing, treatment in a continuous hydraulic classification system, and grinding upgraded a 17% P,O5 reject to 4.5% P,05 with a phosphate recovery of 63%. The CaO: P2O5 ratio was reduced from 2.49 to 1.66.

Jan 1, 1987

By J. Burgess Winter

It is an honor to present the lecture in honor of Daniel C. Jackling, the developer of the technology of copper mining that we continue to use 100 years later. He made a dramatic breakthrough in the early part of this century in the development of bulk-mining methods for large, low-grade porphyry copper-sulfide ore bodies. At Magma Copper, we are particularly proud that the Jackling Award states we have "made significant contributions to technical progress." Before Jackling's time, copper mining had exploited with primitive equipment very high-grade underground ore bodies and high-grade open pits. During the 20th century, the situation changed. Three significant innovations occurred within about 10 years around the turn of the century. In 1903, Jackling's concept of large-scale mining with steam shovels and rail haulage was implemented at the Bingham Canyon mine in Utah. Eighty years after the introduction of rail haulage, the mine converted to truck haulage with large-capacity high-efficiency haul trucks and reduced the stripping ratio from 3.5:1 to less than 1:1.

Jan 1, 1995

By James A. Madonna

Information gathered while studying geological parameters that govern fine gold deposition in the a1luvial environment can be applied to the technological development and selection of fine gold recovery systems. It is recognized that substantial gold losses occur in the minus 1.5 mn (minus 12 mesh) fraction using conventional sluicing systems. Within the past two decades much attention has been given the design and efficiency of these thick bed concentrators. Studies and comparisons drew information from several foreign and domestic placer operations. These studies concentrated on: 1) the physical and chemical properties of alluvial gold 2) the parameters that control fine gold deposition 3) the geometrical form, components and general working structure of conventional sluicing systems 4) the physical properties which contribute to the overall efficiency of a sluice system 5) the comparison of three major styles of conventional sluicing systems. A dramatic improvement regarding fine gold recovery and water quality was observed when using modern multiple-channel sluicing systems in place of conventional single-channel sluicing systems.

Jan 1, 1988

By F. Díaz-Fierros, D. Rubinos, M. T. Barral

The present work explores the reutilization of red mud, an industrial waste arising from the production of alumina, as a component of isolation technologies for hazardous substances. The work involved the study and optimisation of the geotechnical properties of red mud in conjunction with adsorption experiments of As(V) and Hg(II). The results from both types of experiments were integrated in a dispersion-diffusion model to calculate migration curves and minimum thickness of the red mud compacted layer. The geotechnical studies showed that compaction increased the stability and decreased the hydraulic conductivity of red mud barrier. Red mud sorbed high amounts of As (up to 44 mmmol kg-1) and Hg (up to 9.2 mmmol kg-1). The calculated minimum thickness of the red mud barrier was 1-10 cm for As and 2-19 cm for Hg, for an operation period of 35 years. The results obtained let to conclude that the reuse of red mud as a component of isolation systems of hazardous substances represents a viable alternative use of this residue.

Jan 1, 2011

By J. Piper

Markets and Production Beryllium (Be), an extremely light metal, is used in many specialty applications. It the lightest element in Group 2A of the periodic table, with a specific gravity of 1.85 gm/cc, and a high melting point of 12850C (Calvert, 2004). The functional properties of beryllium are stiffness, strength, dimensional stability, resistance to corrosion from acids, and electrical and thermal conductivity. The ability to reflect neutrons and transmit x-rays, makes beryllium irreplaceable in uses such as precision optical components, inertial guidance systems, and in nuclear reactors (Anon, 1993). Beryllium is used in Cu alloys to strengthen and reduce corrosion in copper wire, and in lightweight oxide ceramics that are used as heat sinks, and as electrical insulators. The market revenue of beryllium can be broken down into the seven major groups below, and are plotted on the pie chart in Figure 1 (Anon, 2005). 1.) Telecommunications & Computer 2.) Magnetic & Optical Data Storage 3.) Aerospace and Defense 4.) Automotive Electronics 5.) Industrial Components 6.) Appliance, and 7.) Other.

Jan 1, 2007

By James M. Barker

The occurrence of borates and zeolites at Mesa del Alamo is well documented by McAnulty and Hoffer (1) and Lefond and Barker (2). Further investigation indicates the presence of priceite and meyerhofferite in addition to the howlite (1) and colemanite (2). Analcime occurs in addition to the clinoptilolite and phillipsite (2). The Mesa del Alamo area is about 200 !a south of Tucson, Arizona, and 13 km southeast of Magdalena, Sonora, Mexico, just south of the unpaved road to Cucurpe (Figure 1). The geology and mineralogy of northcentral Sonora is known only on a regional basis except for small areas mapped in detail during copper exploration. The study area is within the Basin and Range province of North America from which borates and zeolites are now produced in California. The borates and zeolites at Mesa del Alamo occur in geological environments similar to those of the Mojave and Death Valley regions of California (3, 4). Thus, a detailed exploration pro- gram in the Magdalena area could result in the discovery of economic borate and/or zeolite deposits.

Jan 1, 1979

By David A. Milligan

This paper describes the Darwin Silver Recovery project, located 66 km (40 miles) outside of Lone Pine, near Darwin, Inyo County, California. A lead-zinc mine and a flotation concentrator were operated there from 1945 to 1957. By the time the mine was shut down, 1.53 Tg (1.69 million short tons) of tailings had been stored in ponds. An average silver assay of 47 mg/kg (1.38 troy oz/st) of tailings stimulated interest for commercial silver recovery. A heap leach process was developed in 1981 at the Tucson Research Center of Anaconda Minerals Company, a division of Atlantic Richfield Company. Agglomeration techniques were used on the ore since the fine grind did not allow direct heap leaching. Lime and cement were required as binders, Early laboratory tests showed that recoveries equal to those from an agitated leach could be achieved at lower capital and operating costs. Laboratory tests determined cyanide requirements, lime and cement usages, and percolation or leach rates, as well as confirming that a zinc precipitation step could recover the silver from the leach solution.

Jan 1, 1983

By R. Mensah-Biney

The kinetics of loading of gold from bromide solution onto an anion exchange resin, PAZ-4, were investigated in a batch stirred reactor. The initial loading rate was described mathematically by a first-order reaction rate equation. The effect of agitation speed, temperature, resin particle size, resin concentration, initial concentration of gold and bromine in solution on the rate was investigated. The model described fairly accurately the loading process in terms of the variables investigated. The rate constant for the initial loading ranged from 0.5 to 1.3 x10-3 sec-1 and the half-time was 9.3 to 56 minutes. The initial rate constant increased with increasing temperature with an activation energy of 15.7 kJ/mole (3.8 kcal/mole). The rate constant was directly proportional to the concentration of the resin in solution. The subsequent reduction reaction was also modeled by the first-order rate equation and the apparent rate constant was estimated to be 3.0 x 10-6 sec-1 with a half-time of about 44 hours.

Jan 1, 1994