Search Documents

By W. de Bruijn

Differential reduction and expansion or shrinkage of hematite under the influence of intimately associated foreign oxides may be a source of stress formation and sinter degradation during blast furnace reduction. To study specific effects, synthetically pre-pared hematite powder evenly doped with 1% by weight of several elements present as oxides, was pressed into tablets and sintered at high temperature prior to reduction. Continuous thermo-gravimetric and dilatometric analyses were per-formed on the sintered tablets as these were being subjected to reduction under simulated blast furnace time/temperature/gas composition conditions up to 1183 K. These findings were coupled with observations on the microstructure of tablets removed from various stages of the reduction process.

Jan 1, 1988

By Staff

GENERAL DESCRIPTION The Lucky Friday mine is located in Mullan, Sho-¬shone County, Idaho. The mining method is horizontal slice cut-and-fill (with timbering for support or un¬timbered) using hydraulic sand. Cemented sand fill at a 1:17 ratio is used in sill pours and at a 1:7 ratio for the final 152 mm (6 in.) capping in successive stope floor pours. Annual production with 100% cut-and-fill mining is 167 829 t (185,000 st). The average depth of mining is 1296 m (4250 ft) and the average travel time to the stope is 45 min. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size The ore body is 396 to 488 m (1300 to 1600 ft) long on strike, 1421 m (4660 ft) on dip, with an average width of 2.3 m (7.5 ft). Dip varies from vertical to1.2 rad (70°) Ground Conditions, Ore Value, and Production Requirements The ore body contains quite strong argentiferous galena and quartz. The hanging wall and footwall are strongly sheared and fractured. Ore value is $211/t ($192/st).

Jan 1, 1982

By M. S. Enders

The McDonald gold deposit is located in the upper Blackfoot Valley in Lewis and Clark County in west-central Montana, about 65 km (40 miles) northwest of Helena and 13 km (8 miles) east of Lincoln (Fig. 1). The McDonald project area covers about 20.7 km2 (8 sq miles) of elevated terrain on the north side of the Blackfoot Valley. The Blackfoot River lies 1 km (0.6 miles) south of the McDonald deposit. East Butte and West Butte rise 300 m (985 ft) above the Blackfoot Valley to an elevation of 1,750 m (5,740 ft). They conveniently divide the area into two parts. The Seven-Up Pete Joint Venture (SPJV) is a partnership between Phelps Dodge Mining (72.25%) and CR Montana, a subsidiary of Canyon Resources (27.75%). SPJV controls about 83 km2 (32 sq miles) of mineral and surface rights in the Lincoln area. The McDonald project is on state-owned and private land. The joint venture has conducted exploration since 1989 at the Seven-Up Pete, McDonald and Keep Cool areas (Fig. 1).

Jan 1, 1995

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 Steve Kral

The former mayor of Los Angeles, CA is a believer in public transportation but thinks neither his city nor the entire country is spending enough on basic infrastructure. In fact, the United States has fallen behind most other developed nations in improving public transportation, the former mayor said. Antonio R. Villaraigosa was the keynote speaker at the North American Tunneling (NAT) conference, held June 22-25 in Los Angeles, CA. The event, held every two years, is sponsored by the Underground Construction Association of SME (UCA). This years? conference attracted 1,091 underground construction professionals from around the world. The accompanying exhibition continues to grow, attracting 142 exhibiting companies occupying 162 booths. As always at NAT, registered attendees received the proceedings volume of the conference, which contains 132 papers. The proceedings volume, including a CD of the papers, is available from SME for $149 member, $139 student member and $189 nonmember. The book can be ordered from SME, NAT item #400-4, phone 303-948-4225, 1-800-763-3132, or online at www.smenet.org/store.

By Bruce Kennedy, Charles Bruce

MINERAL EXPLORATION 'The Function of Exploration Mineral exploration aims at locating the presence of economic deposits and establishing their nature, shape, and grade. As all ore bodies are finite, exploration is essentially the lifeblood of the mining industry. It is divided into two phases, reconnaissance and target investigation. It involves geological mapping, geophysical prospecting (aerial and ground or both), geochemical surveys, prospect drilling, and/or exploratory under-ground workings. Successful exploration is followed by development drilling, mine development, and eventually operation. In underground mining, the term "ore" is applied rather loosely to metalliferous rock that can be profit- ably mined, but nonmetallic industrial minerals and energy related materials constitute a large and growing portion of the tonnage extracted by underground methods each year. It is appropriate, therefore, to consider exploration as the search for any useful substance occurring in the earth or in the oceans that is of sufficient value as to volume, grade, chemical and physical com- position, and location as to make extraction by mining commercially profitable for the investment required. The Selection Process The search for ore includes hundreds of mineral commodities. The more successful exploration organizations tend to restrict the search to the mineral commodities for which they have strong exploration expertise and which have a high probability of being concentrated in areas open to exploration and development in sufficient volume, grade, and proximity to the surface to warrant the large expenditures of risk capital in finding an ore body. Special expertise is needed to engage in mineral exploration and development for the extended periods required for success. Individual prospectors continue to discover mineral occurrences, but it generally takes an organization with strong financial resources to carry the exploration process forward to the point of developing "ore." Still greater corporate resources in the form of management, engineering, and sales capability are needed to develop an ore body into an operating mine. In the past, companies other than those engaged in mining rarely conducted exploration, since they lacked the technical skills and know-how to bring a mine to fruition. A notable trend in recent years has been the success of other resource-oriented firms, such as petroleum companies, that have not only acquired the necessary skills to engage in exploration for hard minerals, but are doing so on a far more intensive scale than known in the industry heretofore. The desire to diversify from increasingly regulated oil and gas, the potential for the discovery of a bonanza, and the large scale of mining projects for these capital-intensive firms are the attractions. US government and certain international organizations, such as the United Nations, have fielded sizable exploration programs in the past decade on behalf of Third World countries wherein many of the world's higher grade deposits and scarce mineral resources are located. The level of overall mineral exploration in the industry is, regrettably, closely geared to the business cycle. Accelerated search during the periods of strong demand, shortages, and high prices, together with the long lead times necessary to develop mineral finds, often creates surges in supply which invariably are followed by a collapse in prices. The trend is, therefore, to exploration funding by larger better-endowed firms having the staying power to weather the cyclical nature of mining. Tax incentives, tariff policies, and other incentives have a decided impact on the level and nature of exploration. Surface mining regulations and environmental forces have fostered a renewed interest in under- ground development despite the lower average costs of surface mining. Spiraling inflation has caused mining costs to outpace prices for many mineral substances. Organizations increasingly are directing their search away from high- grade but limited tonnage occurrences toward low-grade disseminated bulk tonnages amenable to mining on a vast scale. Mining continues, therefore, to grow even more capital-intensive with half-billion-dollar mine-mill- smelter complexes now commonplace, compared with the hundred-million-dollar projects of a decade ago. Reconnaissance Exploration Having elected to search for specific mineral commodities, the exploration company narrows its focus to a reconnaissance of those regions offering the best potential. A working hypothesis which is constantly up- dated concerning the geologic controls for ore occurrence guides the search. The possibilities for discovery are best in areas of existing mines and known ore bodies, where the probabilities are high for additional ore occurrences. In such areas, published data on the trends and geologic controls of known mineral occurrences aid the search. Even in some remote territories, a variety of geologic data is available to the explorationist. The major sources of basic geologic data are the libraries of the US Geological Survey (USGS), the state geological surveys, the American Institute of Mining, Metallurgical, and Petroleum Engineers (AIME), the US Bureau of Mines, and various institutions such as Colorado School of Mines. Overseas, virtually all of the countries have their own geologic survey groups which are associated directly, or indirectly, with their respective ministries of mines. The objective of reconnaissance exploration is to narrow the search by pinpointing target areas having interesting mineralization. These favorable areas are then investigated in detail to delineate potential ore bodies.

Jan 1, 1982

By Tomek Ulatowski

This article discusses four principal areas relevant to financing direct reduction facilities: (1) Lenders' differing requirements with respect to projected cash flows; (2) The impact of leveraging on project viability; (3) The importance of appropriate debt structure; (4) The significance of accurate and complete capital cost estimates. The overall findings emphasize the inseparable connection between the project economics and its financeability, and particularly how direcly intertwined they both are in the project planning.

Jan 1, 1978

By John A. Folinsbee

Placer Dome Canada owns and operates the Campbell mine and mill at Balmertown, in the north western corner of Ontario. The mine operates at 1.3 kt/d (1,500 stpd). It produced more than 9.3 t/a (300,000 oz/year) of gold during the last three years. Changes to the gravity concentration circuit have resulted in a 10% to 15% increase in gravity recovery. The mill head grade at the Campbell Mill is relatively high, averaging 18.5 g/t (0.54 oz/st) over the total reserve tonnage. During the past several years, the head grade was averaging close to 20 g/t (0.6 oz/st). Before 1994, gravity concentration was achieved by using a combination of conventional mineral jigs and tables. Jig concentrates were upgraded on shaking tables to produce a concentrate, which was refined to bullion. In May 1994, management decided to buy two Knelson CD 760-mm (30-in.) concentrators. The objective was to replace the existing jigs and increase gravity gold recovery by up to 15%. Gravity recovery with jigs averaged between 30% to 35%. Gravity recovery was anticipated to increase to about 50% with the Knelsons.

Jan 1, 1997

Residential and commercial plumbers in the United States were introduced to Viega?s copper press technology in 1999. In short, the technology allows plumbers and installers to join copper pipe securely in four to seven seconds on pipe sizes of 1.2 to 10 cm (0.5 to 4 in.) by pressing the connections with specialized tools and fittings rather than through sodering or welding. The water- and air-tight technology, first introduced in Germany in 1989, includes three different sealing elements, a variety of materials and a selection of configurations. By eliminating the need to thread or solder pipe, connections and repairs can be made quickly and securely. Installations are also much quicker and as secure as traditional methods.

Jan 1, 2012

Joy Mining Machinery has developed a new generation of ultra-heavy shields for Anglo Coal Australia?s Moranbah North Mine in the Bowen Basin region of Queensland. These are the largest two-leg longwall shields yet produced by Joy. The longwall at Moranbah North posed a number of challenges, among them a thick, inclined, sometimes undulating seam with increasing depths. The result of this was the need for larger, heavier, more powerful roof supports. Knowing that future longwalls would be going deeper, mine management determined that shield yield ratings of at least 1,750 t (1,930 st) was what would be required.

Jan 1, 2011

By D. Kwak, S. Han, J. Kim, Y. Lee

"INTRODUCTION In carbonate rocks, it is difficult to characterize due to the high degree of heterogeneity in pore geometry. Carbonate rocks have various types of pore systems such as fractures and vugs, and this heterogeneity should be considered for more realistic prediction of oil production in carbonate reservoirs. This study carried out water flooding experiments using five carbonate rock samples in order to observe the change of physical properties. Firstly, qualitative and quantitative analyses were conducted for each carbonate samples by use of XRD and XRF to evaluate the mineral content. All samples are mainly composed of CaCO3 and after that, the ICP experiments were carried out for the water passing through the carbonate. As a result, it was possible to measure the quantity of CaCO3, MgO, Fe2O3 component in water and due to this effect, the permeability of carbonate was altered especially in the front part of the core. Finally, after the carbonate was saturated by 30 API crude oil, water injection to enhance the oil recovery was repeated and it was confirmed that the oil-producing characteristics are different depending on the dissolution characteristics of carbonate rock samples. CORE SAMPLE ANALYSIS Porosity and Permeability Experiment Porosity and permeability were measured. The procedures are as follows (Figure 1). 1. Measure the diameter and length of the dry core using Vernier caliper. 2. Put the core in the core holder, run vacuum pump until the inlet and outlet pressure gauge equal. 3. Saturate the core using ISCO pump to inject water at 650 psi 4. Measure the amount of water and calculate the porosity 5. Set up the outlet pressure at 500 psi. 6. The flow rate measure when the equilibrium 7. Calculate the permeability through Darcy’s Law. 8. Cutting of core upstream and downstream about 13cm, and measure the each of permeability."

Jan 1, 2016

By H. M. Woodrooffe

In the U.S.S.R. three of the world's major-known phosphate deposits are located. These have an estimated reserve of 2,600 million short tons of contained phosphorus. The best known lies in the "Khibiny Massif', an alkalaic rock formation some 600 square miles in area with mountain peaks rising to an altitude of 3300 ft above the surrounding terrain. Location and Climate The Khibiny Massif lies north of the Arctic Circle in the Kola Penin-sula, 100 miles south southeast of Murmansk, at a latitude of 67° 30' north and a longitude of 35° 30' east. It is serviced by rail from Murmansk and Leningrad (900 rail miles south). Figure 1. The phosphate operations are centered about the two cities of Apatity and Kirovsk. The former has a population of 35,000 and is on the Leningrad¬Murmansk railway. Kirovsk, 15 miles to the east, is now a city of 85,000. The region is at the limit of the tree line with most of the ground cover typical tundra. The mean annual temperature is -4°C and, in a normal year, precipitation occurs on 240 days. Because the temperature does not exceed freezing for 300 days each year, virtually all of the precipitation is snow. The two towns are nestled between mountains and, because of the weather, extensive avalanche control systems incorporating major engineering construction have been introduced to divert snow slides away from the populated areas. General Kirovsk was first settled in 1923 and the apatite occurrences were recognized shortly after. In 1933, the first concentrator for the recovery of the mineral was completed, producing in its first year of operation 170,000 tons of concentrate. With the growth and demand for phosphate in the U.S.S.R., a second concentrator was erected at Apatity. It calve on stream in 1963. Because of the severe climate, in the early days mining was limited to 85 days per year. With the introduction of new techniques and improved mining methods, the season has been extended to 115 days.

Jan 1, 1972

By E. Bonates

Dispatching systems have attracted considerable attention in the last years due to substantial gains in productivity achieved through their implementation. These systems will be extensively used in the future by mine operators as a means of optimizing productivity. This paper describes a computer simulation model designed to study the feasibility of continuously dispatching trucks in open-pit mines based on different operating procedures. The system developed is semi-automated and performs an active role by suggesting to the dispatcher the best assignment according to each specific circumstance. It is modeled based on an advance-clock increment to reduce computing time and allows the insertion of heuristic dispatching policies and equipment real-locations when breakdowns occur. Dispatching procedures taking into account long-term production objectives are considered in system development.

Jan 1, 1988

By W. M. Houston

Pebble phosphate deposits of Florida have been established from south of Hardee County to north of the Georgia state line or a distance of over 200 miles and mining has been carried out in an ever-increasing scale since 1885. Deposits that can be mined on the economical standards of today, however, are confined in a roughly circular area about 30 miles in diameter, generally within Polk and Hillsborough Counties, in the central part of the State. The phosphate deposit, which is in the Bone Valley formation of the Pliocene age, is covered by Pleistocene terrace deposits of unconsolidated quartz sand. The matrix is composed of approximately 1/3 clay slimes, 1/3 silica sand and 1/3 phosphate pebbles. The phosphate occurs usually as rounded pebbles, ranging in size from about 1 inch down to approximately 400 mesh. The phosphate may vary from well- defined hard pebbles to soft weathered rock. The color and character of the phosphate will range from smooth buff pebbles to black and white or "salt and pepper rock." The screen size distribution varies considerably between locations and the recoverable phosphate varies may extend down to the 400 mesh size in some areas, while in others the phosphate may cut off sharply at about 35 mesh.

Jan 1, 1962

By Larry D. Patts

Section 317(e) of the Federal Coal Mine Health & Safety Act of 1969 directed the Secretary of the Interior to prepare standards under which all working places in a mine shall be illuminated by permissible' lighting while persons are working in such places. Section 317(e) further provides that such Working places shall be illuminated within 18 months after such standards are promulgated. In accordance with this section of the Act, there was published in the Federal Register for December 31,1970 a notice of proposed ru1emaking which prescribed the illumination to be provided in the working p1 aces of underground coal mines. In 1ight of written comments, suggestions, and objections to this proposed ru1emaking, the proposed standards were with-drawn and reproposed in the Federal Register for Wednesday, October 27, 1971. In light of further comments, suggestions, and objections, a public hearing was held on April 4, 1974 and standards were again reproposed and published in the Federal Register for April 1, 1976. Promulgation of the final lighting standards took place on October 1, 1976 which means that the underground coal mining industry must comply with face illumination requirements by April 1, 1978. As mentioned previously, the first proposed ru1emaking for illumination of underground coal mines was published in the Federal Register on October 27, 1971. In early 1972, Consolidation Coal Company (Consol) and the United States Bureau of Mines agreed to a cooperative study of underground face illumination; Conso1 felt that expertise in this field would become increasingly important. Conso1's initial efforts in illumination were aimed at investigati.ng practical lighting systems for underground face equipment. We were concerned with installing unobtrusive lights which provided sufficient face illumination for safety, but at the same time were readily maintainable, electrically reliable, and physically sheltered from damage.

Jan 1, 1977

Potassic alteration domains of greenstone belt lode gold deposits are characterized by systematic partitioning between different groups of incompatible elements, such that Al, Ga, Th, U, Ti and V are decoupled from K, Rb, Ba, Cs and Tl. The former group of elements have concentrations and interelement ratios in alteration domains which reflect host rock control, implying approximately isochemical behaviour. In contrast, K, Rb, and Ba are strongly coenriched, and linearly correlated over th4ee orders of magnitude in abundance, where K/Rb = 260, and K/Ba = 40: K/Cs (-10 ) and K/T1 (-2x10 ) are also correlated, though more weakly. Lithium abundances, and Rb/Sr ratios are erratic in altered rocks. These interelement trends, collectively, are present in deposits variously hosted by ultramafic, mafic, or felsic volcanic rocks, and sediments or granitoids.

Jan 1, 1988

By Joaquin Ruiz

Although Mexico is known for its important base-and precious-metal deposits, and lately oil, some industrial minerals play an important role in the Mexican economy. Sulfur, the most important industrial mineral produced, accounts for approximately 7.5 percent of the total value of mineral production. This commodity is followed by fluorspar, which accounts for 6.4 percent. Predominantly, the fluorspar deposits are very high grade (75 percent CaF) and are contained in lower Cretaceous limestone in the states of Coahuila and San Luis Potosi. Base-metal deposits in the area of Parral (notably San Francisco del Oro), however, have become important fluorite producers in their own right. (San Francisco del Oro has 12 percent fluorite gangue.) Salt from the Guerrero Negro solar-evaporation ponds in Baja California account for 6 percent of the total shares; phosphate from Baja California also accounts for 6 percent. Barite from man to deposits in Lower Cretaceous limestone in Coahuila and exhalative deposits in the Paleozoic section of Sonora account for 1.0 percent. In addition, Mexico is an important producer of gypsum, graphite, and celestite. Although graphite and celestite are not large-scale operations and thus do not contribute much to the Mexican economy, the Mexican graphite and celestite deposits are among the most important in the world; Mexico is one of the world's few sources of celestite. The celestite is in the form of large, high-grade mantos in Cretaceous limestones in northern Mexico.

Jan 1, 1987

By Roe-Hoan Yoon, Kaiwu Huang

Derjaguin and Dukhin [1] were the first to develop a first-principle flotation model by considering the role of surface forces in bubble-particle interactions. According to the model, the ζ-potentials of particles constitute the energy barriers, E12, for particle (1)-bubble (2) interactions, which corroborates well with the observation that flotation recoveries reach maxima under conditions of zero-zeta potentials. In the present work, a series of surface force measurements were conducted for the quartz-dodecylamine hydrochloride (DAH) flotation system, for which detailed studies had been conducted by D.W. Fuerstenau [2], including the measurements of zeta-potentials, contact angles and collector adsorption as a function of pH.

By John W. Amos, John A. Christophersen

The Deister No. 7, "77" and "88" Coal Washing Tables have been used by the coal industry in the cleaning of finer sizes of coal (-1/2") since washing of coals became necessary. This paper will deal with the in-plant use of the CONCENCO "88" Deister Coal Washing Table, along with a brief description of the Deister Refuse Superscalper. Efficient in-plant operation of the Deister Coal Washing Table can be accomplished by the adjustment of six basic variables. These variables are: 1. Headmotion 2. Feed 3. Riffle Pattern 4. Dressing Water 5. End Elevation Adjustment 6. Side Tilt Adjustment VARIABLE NO. 1 HEADMOTION The reciprocating movement of the "88" Table is developed within the headmotion entirely by rotating eccentric weights. Four parallel shafts each carry pairs of eccentric weights (Figure 1-A). The shafts are geared together with gears of 2 to 1 and 1 to 1 ratios, so that the two shafts in vertical alignment which carry the small eccentric weights operate at twice the speed of the other two shafts. The other two shafts are also in vertical alignment with each other and carry large eccentric weights. As power is supplied to one high-speed shaft, once each revolution the forces produced by the large eccentric weights are reinforced and then opposed by the small eccentric weights in a horizontal plane. The direction of movement of the table decks is reversed much more sharply at the forward end than at the backboard end of the stroke, causing material to move toward the refuse end of the table. Because in the vertical plane all centrifugal forces oppose each other and cancel out, there is no vertical movement and virtually no impact to the support steel. The "88" Coal Washing Table is suspended by wire-rope cables from three overhead steel beams, and operates with a perfectly uniform and exact differential motion curve in every stroke (Figure 1-B). The decks, being attached to the headmotion, are reciprocated in a parallel line from the head- motion to the refuse corner at speeds varying from 260-300 strokes per minute. Both length of stroke and speed are set at the factory for the coal characteristic desired and should not be changed. A 3 HP motor is mounted on the top of the headmotion and uses 3 V-belts to transmit driving power to the headmotion. The headmotion must be kept running on a horizontal plane. Punch marks are provided on the side and end of the main frame for use in leveling the motion (Figure 1-C). In leveling the motion, the work must always be done with two cables. To level the long side of the headmotion both front or both rear cables are raised or lowered to bring motion back to level. To level side to side, both right side cables or left side cables are raised or lowered to bring motion back to level VARlABLE NO. 2 FEED When tabling different size coals, each deck of a "88" DIAGONAL-DECK Coal Washing Table will handle feeds as described in figure 2-A.

Jan 1, 1988

By J. Gonzales

For this paper, blending is conceived as a linear process that takes a proportion from one stockpile or location and mixes it with another proportion from another stockpile or location, with the aim of obtaining a set of attributes in the final combination that satisfies a given quality standard. A blending class is the combination of ore attributes and grade limits that should be stockpiled in order to achieve the target quality for the process Plant. Current literature considers the number of blending classes in the Open Pit Scheduling Problem as fixed; however, this paper shows how to define specific blending classes for a given deposit so that the schedule can be solved with Sequential Linear Programming. Finally, a practical application in a Peruvian copper mine is shown, involving four ore attributes: alteration code, arsenic, oxide content and copper grade. INTRODUCTION In copper ore deposits, blending for quality has not been a typical issue as it was for coal and iron ore; however, current copper deposits in the world are becoming geologically and chemically more complex than in the past and is not uncommon to have quality standards involving up to four attributes, introducing an additional complexity to the mine scheduling process [1]. In order to solve the Open Pit Scheduling Problem with multiple ore quality constraints, two assumptions are necessary [2]: - Grade constraints are defined only by specifying an upper and lower limit; - Ore attributes in stockpiles are pre-determined, i.e., the upper and/or lower grade limits defining each stockpile are already known at the time of setting the schedule. Assuming each ore attribute is classified as high, medium or low, and there are four attributes to be controlled in the target ore quality, the number of possible blending classes would be 3^4 = 81. This number is impractical to manage in Operations, and increases the complexity at the time of running the schedule; therefore, a more efficient way of classifying ore is required. Note that in practice, one blending class can be split into more than one stockpile location due to physical capacity constraints. Therefore, the number of stockpiles could even be more than 81. BLENDING CLASSES DETERMINATION For a given set of ore grades with lower and upper limits to be achieved, one has to find all feasible ore classes such that when blended, their grades will lie within those limits.

Jan 1, 2019