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By Flávio de Castro Dutra

Este artigo apresenta a influência da temperatura e da atmosfera de secagem sobre a resistência mecânica de um aglomerado esférico, com diâmetro entre 3 mm e 8 mm, constituído de pellet feed e silicato de sódio, produzido através de pelotamento em discos e endurecido em temperaturas até 550ºC. O aglomerado obtido a partir deste processo apresenta características que permitem que o mesmo possa ser transportado a longas distâncias e manuseado até a sua utilização na sinterização. As principais técnicas experimentais utilizadas neste estudo foram difração de raios-X, espectroscopia Raman, microscopia eletrônica de varredura e análise térmica. O aglomerado foi submetido à secagem em temperaturas entre 100°C e 550ºC e atmosferas constituídas de ar atmosférico e/ou CO2. Durante o processo de secagem, observou-se a influência da temperatura e da atmosfera sobre a resistência à compressão do aglomerado após exposição à água, assim como na morfologia do silicato de sódio, evidenciando uma dependência entre elas. O uso de CO2 durante a secagem mostrou-se eficiente para a formação de um aglomerado com maior resistência mecânica após exposição à água.

Aug 2, 2018

By H. T. Fargey

This paper considers the social, economic and political factors and their influence on future production and marketing of Western Canadian minerals. Among the subjects discussed are inflation, taxation, anti-combines legislation, unstable international currency relationships, economic nationalism and protectionism, the European Common Market and environmentalism.

Jan 1, 1973

By WILLIAM COUPER

THE completion of the excavation of the tunnels through Bergen Hill for the extension of the Pennsylvania Railroad into New York City, together with the fact that the excavation for the rest of the tunnel-line has been practically completed for some time past, arouses sufficient interest to justify a description of the engineering- and contracting-work, and the organization by which it has been accomplished. The finishing of the excavation is a long step towards the completion of the entire tunnel-project. It will be interesting to the members of the Institute, who last spring were taken through a great part of this work, to know that by the excavation now completed it is possible to walk by tunnel from a, point near Sunnyside yard, Long Island, east of the 34th Street ferry, to the central passenger station site in the City of New York, and thence by tunnel to the Hackensack meadows, and by an embankment constructed across these meadows to Newark, N. J. This New York tunnel extension is being constructed as a part of the Pennsylvania Railroad system by the Pennsylvania Tunnel & Terminal R. R. Co., a corporation of the States of New Jersey and New York. It extends from a point near Newark, N. J., to the Sunnyside yard, Long Island, as indicated on the accompanying plan.

Jan 7, 1908

By AIME AIME

WAR undoubtedly accelerates metallurgical progress, although its most obvious effect is a tremendous waste of materials. The necessity for restrictions in normal uses of metals results in a search for substitutes or for enhanced properties that will make a limited supply last longer. An insatiable demand for more metal to be produced more quickly favors trial of new processes, new metallurgical tools, and new methods of fabrication. Greatest advance is made, of course, in practical physical and process metallurgy. Theoretical or fundamental research necessarily must take a secondary position, and is so regarded in this review. It is not that developments in theoretical physical metallurgy are lacking entirely, but, following the same trend noted in other countries as they entered war activities, attention in the U.S.A. has been shifted strongly to those developments having an immediate practical application.

Jan 1, 1943

By W. P. Haynes

To evaluate the potential use of battery-powered vehicles, the Bureau of Mines studied the performance of a battery-powered truck propelled by a 9.6 kw dc motor in considerable detail under various driving conditions. Some of the factors investigated were: Gear ratios, payload (1,468 lb maximum), road grade, stops made per mile, and charging time. With the exception of some difficulty with brush wear, operation and maintenance of the truck was generally trouble-free. On range tests carrying an 868-lb payload a distance of about 50 miles at an average speed of about 19 mph in normal driving, the Bureau's battery-powered truck averaged 1.6 to 1.9 miles per ac kwhr input. Stop-and-go tests indicated that prospects for the use of battery-powered trucks in house delivery service within city limits are good; performance of the Bureau truck was sufficiently high for that type of service. The estimated power cost of 2.65 cents per mile while making 10 stops per mile was attractively low from the standpoint of fuel costs.

Jan 1, 1967

By John J. Stetler

Introduction Samarco Mineracao S.A. is a Brazilian corporation jointly owned by S.A. Mineracao da Trindade (SAMITRI) and Utah International Inc., with 51% and 49% interests respectively. Headquartered in Belo Horizonte in the state of Minas Gerais, Samarco owns and operates an open-pit mine capable of producing 10 Mt/a (11 million stpy) of iron ore; a concentrator rated at 7 Mt (7.7 million st); a 12-Mt (13.2-million st) capacity pipeline transportation system; a 7 Mt (7.7 million st) filtering and pelletizing plant; and a harbor facility. The mine, concentrator, and initial pipeline pump station are located at Germano, near the town of Mariana, about 300 km (186 mi) north of Rio de Janeiro by air. The pipeline terminal, pellet plant, and harbor are at Point Ubu, 70 km (43 mi) south of Vitória the capital city of Espirito Santo State, and about 300 km (186 mi) by air east of the mine. The 396-km (246-mi) pipeline uses 508-mm (20-in.) pipe except for a 53-km (33-mi) section where 457-mm (18-in.) pipe is used to absorb and control slack flow through increased velocity in one of the areas along the route. Iron Ore Slurry The hematite iron ore concentrate produced at the Germano plant is processed by grinding the average 53% run-of-mine itabirite iron ore to -150 µm (100 mesh) for gangue liberation. It is treated further by cyclones to remove the 10 µm (1250 mesh) slimes to prepare the ore for flotation. During the multistage cationic flotation operation, the silica is reduced to make low silica (-0.9%) concentrate for the direct reduction (DR) iron-making process or to make normal silica (-3.5% +2.5%) concentrate for the blast furnace (BF) iron-making process. The concentrate then is reground further to meet the particle size requirements for pelletizing. The reground hematite concentrate is screened in closed circuit with the regrind mills for close control of the maximum particle size, not to exceed 3.5% +75 µm (200 mesh) for pipeline wear protection. Thickened concentrates at 70% solids are pumped to four 5-kt (5,510-st)(dry) agitated slurry storage tanks that hold the feed before it goes to the first pipeline system pumping station. Pipeline Capacity The pipeline system was designed to transport 12 Mt/a (13.2 million stpy) at 65% solids with an average minimum safe operating velocity of 1.5 m/s (5 fps). The pipeline will transport 5 Mt (5.5 million st) in 1982. The system, therefore, is moving the slurry in 13-kt (14,330-st) batches, dispatched at intervals to move 20 kt/d (22,050 stpd). In operation, the pipeline has exceeded its 93% availability design factor. Pipeline Facilities Pump station No. 1 is located at the Germano minesite 1 km (3,330 ft) above sea level. It has the following principal equipment: • four 5 kt (5,510-st) (dry) agitate slurry storage tanks; • two centrifugal charge pumps (one stand-by); • a 217-m (712-ft) safety loop; • seven mainline positive displacement triplex plunger pumps (189 m3/h, or 832 gpm, at 14.9 MPa, or 2,162 psi, maximum output pressure); • four corrosion-erosion spools; • a control and command center. Operation The output at pump station No. 1, with 66% solids, is pumped 151

Jan 10, 1982

By F. L. Robertson, C. H. Bacon, J. W. Till

A description is given of the development of a process for making low sulphur steel on furnaces fired with 2 1/2 pet sulphur oil. Slag and metal are analyzed at melt. Slag weight is estimated. A simple method for calculating necessary lime additions for making 0.025 and 0.035 pct sulphur steel is given. DURING the last 15 years John Summers & Sons have been confronted increasingly by the problem of producing a low sulphur steel from high sulphur materials. The principles adopted, the processes used, the data collected on a furnace specially set aside to study this problem, and speculation on future developments are dealt with in this paper. This study is limited to two melting shops which make sheet steel only, of the following analysis: C, 0.06 pct; Mn, 0.33; S, 0.04, 0.035, 0.03, and 0.025; and P, 0.02 and 0.015. All the processes described have reference to the manufacturing of this steel and are not necessarily applicable, without modification, to other steels. The raw materials used are given in Table I. Development of Process About 1934, Schenck's work on phosphorus was applied to frame quantitative rules for the removal of phosphorus, with great success. We could however, neither in our own work nor in the voluminous literature, find any assistance to frame quantitative furnace rules to remove sulphur. Perhaps, we should have persevered with Schenck's outstanding chapter on sulphur in The Physical Chemistry of Steelmaking, but we missed the importance of one (Ss) sentence on p. 478. "Evidently this ratio -------- is [S] the important quantity to be considered from the point of view of chemical reaction; it is clear that the sulphur content of the steel alone, as frequently given in the literature cannot suffice for judging the success of the reaction unless special details are given regarding the sulphur charge and the amount of slag."' We were not however very successful in getting values for Schenck's "n" that agreed with (S) arrived at by our direct analysis. [S] In addition to Schenck's work much attention was given to V value. The formula selected for V was: (SCaO) — 1.57 (2 P2O5) (S SiO2) After plotting many charges a curve was obtained, based upon the relationship between the percentage sulphur in the finished steel and V value, Fig. 1. For a long time this was used as standard with about a 60 to 70 pct success on producer-gas fired furnaces, if all uprising scrap was used and no bought scrap used. An inspection of this curve shows clearly that it is not a quantitative manufacturing process, and that it was not sure enough to go to the trouble and expense of doing stage analysis for P,O,,

Jan 1, 1952

By B. R. MacKay

Introduction Of all Canada's mineral resources, its coal deposits have probably received the most continued attention by the Geological Survey since its inception ninety years ago. A casual review of the many geological maps of coal areas by the Geological Survey will show how geological and topographical detail on the maps has gradually increased to meet the ever urgent demands of the mining industry. In the early stages of exploration, the mapping done, especially in the coalfields of the Rocky mountains, was largely of a reconnaissance nature. On these maps the topographical relief, where shown, is indicated by form lines or by hachures, the boundaries of the coal-bearing formations are only approximately defined, and the geological structure of the coal seams is illustrated by one or more generalized cross-sections. Once mining development on a coal area began, there arose a demand for more detailed maps. To meet this need, topographical and geological maps on a scale of one inch to a mile were made, in which the ?topographical relief is shown by contour lines having generally an interval of 100 or 200feet. On the geological maps, the boundaries of the formations are accurately delineated, and the geological structure of the area is illustrated by one or more structure sections drawn across the area, which show the position and attitude of fault lines, the crests and troughs of folds, and the attitude of the various formations (See maps of Costigan Coal Field, 1905, and Cascade Coal Basin, 1907, by D. B. Dowling).

Jan 1, 1932

By R. Scheffel, R. Washnock, G. Zarate

"The second survey of global copper leaching operations has been completed, the data published in 1998 has been updated. A survey questionnaire was sent to 40 identified copper leach operations around the world including those leach properties still in operation that responded to the 1998 survey. This paper lists the results of 26 operations that responded to the survey. Selected information from the survey is represented in graphical form. INTRODUCTION Copper leaching was first performed by truck dumping huge monolithic stockpiles of ore and flooding the surface with leach solution or lixiviant with little understanding to the parameters that support optimum recovery of the metal. Heap leaching has evolved into a controlled process whereby a complex low grade ore is stacked in shallow lifts (under 10 m in most cases). The ore is usually crushed and agglomerated, conveyor stacked on an engineered containment system (leach pad) and irrigated in a controlled manner. The lixiviant chemistry is designed to extract the optimum amount of metal from the ore. At first glance, heap leaching appears to be simple in nature. In fact, it is a very complex process. A plethora of technologies such as material handling, environmental control, analytical techniques, solution chemistry, hydrology and many others contribute to the design and operation of a heap leach. Heap leaching requires an inter-disciplinary approach between geologists, mining engineers and metallurgists. Many operational factors affect heap leaching. The process must address climatic factors, flexible operation, quality control before and after leaching, physical characteristics of the ore, geology, mineralogy, reagent consumption, solution application rate and methods (wobblers or emitters), particle size, and favorable conditions for bacterial growth and survival. Predicting, troubleshooting, and optimization of leach recovery is a significant challenge when considering the number of variables involved and their interactions The survey updates many of these key operating parameters to the current state of the art. It should provide the reader an excellent benchmarking summary of these factors on a site-by-site basis."

Jan 1, 2016

By G. M. Kintz, E. H. Denny

Although some engineering colleges recognize the importance of instructing young engineers in industrial safety , the different institutions handle this instruction in various ways . It is believed that a recent industrial safety course , conducted at the Colorado School of Mines through the cooperation of this school and the Bureau of Mines , U. S. Department of Commerce , presents some unusual features that are of interest to other schools and to the mining public .

Sep 1, 1930

By A. M. Gaudin

The study that is presented here was undertaken in order to condense information concerning comminution, and covering a great variety of conditions, into one or several rules which would be of use in the development of a systematic theory as well as in many directly practical ways. As the result of this work, the following conclusions appear justified: 1. Rocks may be classified, from the point of view of comminution, into two general classes which are (a) structurally homogeneous, and (b) structurally heterogenous. Structurally homogeneous rocks are those in which fracture takes place through grains and grain boundaries alike; heterogeneous rocks are those in which fracture passes in preference through grain boundaries. 2. If a sized product composed of structurally homogeneous rock grains is crushed, a product results in which the relation between the percentage weight of grains of various sizes and these sizes themselves follows a definite law. This is not true of structurally heterogeneous rocks. 3. Ball milling with structurally homogeneous rocks, and under certain conditions, presents exceptions to this rule. 4. In ball milling with structurally homogeneous rocks, there exists a critical ratio of size of feed to size of balls, above which the machine operates essentially by surface wear, and below which it operates essentially by breaking across the body of the particle. 5. In ball milling, with structurally homogeneous rocks, and when the size of feed grains exceeds the critical size, the particles worn off are of a substantially definite size which depends upon the ball size. 6. In the case of a typical heterogeneous rock, such as Manhattan schist, particles are crushed in such a way that there is a greater proportion of grains of the size that is the average grain-size of the rock than there would be in that range of size with a structurally homogeneous rock. This has been found to be true with all the crushers tested. 7. Different methods of crushing produce grains of different shapes. 8. The importance of sizing the — 200-mesh portion of a crushed product, in crushing-efficiency investigations, has been greatly under-

Jan 1, 1926

By Charles A. Anderson

Arizona and western New Mexico contain 17 of the 25 leading copper mines in the United States. Production of molybdenite, lead, zinc, and by-product gold and silver is important. Precambrian ore deposits are largely limited to the Yavapai Series in central Arizona and most of the metal production has come from massive sulfide deposits, dominantly pyritic but containing different amounts of chalcopyrite, tennantite, sphalerite, and galena. Their origin has been interpreted by some geologists as hydrothermal replacement of schist, and by others as syngenetic deposits later metamorphosed. Most of the important ore deposits are of Mesozoic and early Tertiary age, dominated by the porphyry copper deposits spatially related to stocks, plugs, sills, or dikes of quartzbearing porphyritic to granular rocks. Host rocks for these deposits include the associated intrusive rocks, Precambrian schist and granitic rocks, Paleozoic limestones, Mesozoic(?) porphyritic andesite, and Mesozoic arkose and siltstone. Chalcopyrite and pyrite are the dominant hypogene sulfide minerals in the porphyry copper deposits, accompanied by minor molybdenite. The Questa deposit in New Mexico is largely molybdenite. Supergene sulfide minerals are chalcocite and minor covellite. Some ore bodies consist largely of hypogene sulfides, some are dominantly supergene sulfides, and others are mixed hypogene-supergene sulfides. Supergene oxide-copper minerals locally are important. Intense hydrothermal alteration accompanied the metallization in all of the porphyry copper deposits; five types are recognized, ( 1) propylitic, (2) argillic, (3) potassic, ( 4) quartz-sericite (without clay), and (5) lime silicate. The origin of the closely spaced fractures that contain ore minerals in intersecting veinlets is debatable; some geologists favor tectonic forces and others suggest cooling of magma or shrinkage due to alteration. Breccia pipes, veins, and limestone replacements account for some important deposits of Mesozoic and early Tertiary age, and presumably these are genetically related to the various intrusive rocks of this age. The St. Anthony deposit of Cenozoic age in southeastern Arizona is unique, consisting of a lead-zinc vein associated with later molybdenum and vanadium oxide-minerals. Cenozoic manganese deposits in western Arizona are an important resource in the Artillery Mountains, but the extent to which they may be exploited is a metallurgical and economic problem.

Jan 1, 1968

By Robert A. Jankowski, Charles Babbit, Natesa I. Jayaraman

The Bureau of Mines, together with Foster Miller Inc., has developed a compact, high pressure scrubber system for continuous miners. Powered by water energy alone, it can move up to 1.4 m3/s (3,000 cfm) of air and scrub a significant amount of the dust in the air. Full- scale model and underground trials showed excellent performance. Operated at 8.3 x 10[3] kPa (1,200 psi) and 38 L/m (10 gpm), the scrubber has reduced operator and return dust concentrations by more than 50 pct. This paper describes the full-scale model development, system fabrication and underground evaluations of this novel and inexpensive scrubber.

Jan 1, 1989

By A. Acar, E. T. Tuzcu, Y. O. Vural

"In order to sustain the operations in chromite processing plants, due to concentrate selling cost, plants are in need of optimization in terms of capacity and recovery increase. In this study, DAMA Engineering carried out a methodological study for the purpose after a detailed sampling campaign in a plant which is situated in one the main chromite zones in Turkey; 1 The current status of the plant was investigated quantitatively with the bottlenecks, Plant overall efficiency in terms of recovery, Efficiencies of the particular equipment (mill, screen, gravity circuit), 2 A new circuit was proposed with the available equipment, 3 The new circuit that is product of modeling and experimental effort, was evaluated quantitatively as if it is a running plant, After 4-5 months continuous effort, the capacity of the plant was almost doubled (37 tph to 70 tph) and the plant overall recovery was increased from 48% to 63% by grinding adjustments and additional shaking tables theoretically. It was shown that the plant may move into profit with the minimum amount investment using readily available system. INTRODUCTION TO THE CURRENT PLANT The objective of the study is to increase the capacity and recovery of the chromite plant owned by Kurmel Group, Akmetal Inc. The complete study is given elsewhere (DAMA, 2016). The plant is situated in one of the main chromite zones in Turkey; Aladag. The current capacity of the plant is around 35 tph and the overall recovery declared by plant engineer is around 47 %. The chromite plant mainly has 3 units; crushing, grinding and gravitational circuits. The crushing circuit has 3 stages; a jaw crusher, a secondary cone crusher and a tertiary cone crusher operating in closed circuit with a vibrating screen with the aperture size of 15mm. In grinding circuit, a rod mill (2.44 6m x 3.6) and a ball mill (2.13 x 2.3m) are used for primary and secondary grinding, respectively. Derrick screens are used for the size separation. Additionally, for the size classification of the system, hydrosizers are used and each size fraction is sent to the shaking table sets. The middling product of the tables is subjected to re-grinding with a ball mill with the dimensions of 2.1 x 2.9 m. The capacity and the recovery are affected by several bottle necks; hydrosizers’ pure classification, primary/secondary mill power vs capacity problems, re-grinding product’s high slime content and so on. These combined factors made the plant to run in low recovery and capacity. Hence, DAMA was asked to make a study to investigate current status of the plant and propose a new solution to increase capacity and recovery of the plant. The conceptual flowsheet of the circuit is given in Figure 1 below."

Jan 1, 2017

? Hollow Stem Auger ? Drill Bits ? Leads/Torque Arm ? Hydraulic Gearbox ? Hydraulic Power Unit ? Grout Pump ? Crawler Crane

Jan 1, 2002

By C A. Chaves

During most of its history, the metallurgical industry has been seen as a source of pollution for the environment. However, during the second half of the 20th century, as the world environmental legislation became more protective and the global environmental conscience consolidated, this landscape began to change. Metallurgical activities have become increasingly related to the social and environmental impacts caused by human activities. With the adoption of environmental management in steel plants, there has been a significant change in steel-related activities which have begun to introduce modern plants, managed with environmental, social and productive efficiency. More and more, steel companies search for a decrease in negative impacts of their industrial operations on the environment, fitting them to the current environmental legislation and even increasing productivity and profitability. The steel industry has started to become a clean and more efficient industry. This work presents a case study of a sustainable steel plant project, located in an environmental preservation area. The project includes treatment and reuse of industrial liquid effluents, air emissions control, noise control, recycling, recovery and adequate waste management. It also incorporates production processes using clean technologies, capable of coproducing energy resources which may be used in the plant showing how is it possible to produce steel with environmental and social responsibility and even increase the enterpriseÆs profitability.

Jan 1, 2009

By J. M. Kurtz

Ten Methane Monitoring Systems and Machine Control Circuits were examined and tested per the requirements in CFR Title 30, Parts 18 and 27, the Standard UL 913 and the document "Test Requirements For Instruments or Apparatus to be Considered for M.E.S.A. Intrinsically Safe Certification (Tentative)". The test results indicate that none of the products complied with all parts of all requirements as interpreted by Underwriters Laboratories Inc.

Jan 1, 1980

One of the largest mining shows in the world is set take place April 15-18 in Santiago, Chile. Expomin 2008 is expected to attract more than 60,000 professionals from around the world who are involved in all aspects of the mining industry. Expomin has become an important place for mining leaders to meet, organizers said, ?with great expectations for the development of new business in international markets.? This is the 10th Expomin. The event, held every two years, is expected to at-tract exhibitors from more than 25 countries, representing nearly 1,000 companies and more than 3,000 brands related to the mining industry, organizers said. Most all of the world?s largest equipment manufacturers and suppliers ? surface and underground ? will be represented. The importance of the event to the Chilean and international mining industry can be seen by the increase in size of the exhibition ? 13 percent more surface space in 2008, compared with the 2006 Expomin. The exhibit will be broken down into four pavilions, covering 34,054 m2 (355,570 sq ft). An outdoor area for the large mining equipment will cover 16,120 m2 (173,520 sq ft).

Jan 1, 2008

By Thomas D. Henderson, Wayne R. Hopkins, Arthur Lynch

Brought on stream in 1975, the 10,000-tpd copper oxide processing plant of Anamax Mining Co. at Twin Buttes, Ariz., is now at design capacity. Based on acid agitation leaching, solvent ion exchange, and electrowinning, the Anamax plant bridges an ore to metal gap by producing mine site copper cathodes from lime-bearing oxidized feed. Until recently, such material was considered submarginal because copper recovery was uneconomic using conventional methods. The new Twin Buttes oxide plant, the largest of its kind in North America, is one stage in the expansion of Anamax, a joint venture owned equally by The Anaconda Co. and Amax Inc. Other phases include an accelerated stripping program in the pit, an expansion of the sulfide concentrator from 32,000 to 40,000 tpd, and construction of a 57 MW diesel generating station for the oxide plant. Once the basic oxide flowsheet had been selected, Anamax awarded a contract to a joint venture of Arthur G. McKee and Co.' (Western Knapp Engineering Div.) and Davy Powergas' (a member of the Davy International group). The contract covered process design, engi¬neering, procurement, and construction. The property, identified by Banner Mining Co. in the 1950s and developed by Anaconda, was put into production in 1965 as an open pit mine equipped with a 32,000-tpd sulfide flotation concentrator. The Twin Buttes mine has since become one of the largest earth¬moving operations in the world, excavating some 10 tons of waste per ton of ore. At the original sulfide mill feed rate, the mining fleet moved 375,000 tpd of material (225,000 tpd alluvium and 150,000 tpd of rock). The most significant copper mineral in the ore body is chalcopyrite, which is scattered in pockets of weak mineralization. Cap oxidation has progressed to depths of several hundred feet, principally as chryso¬colla. Present estimates place minable reserves at 347 million tons of sulfide ore grading 0.6% copper and 55 million tons of oxide ore grading 0.9% recoverable copper from 1.25% total copper. The nonre¬coverable fraction is composed of native copper and insoluble copper oxides and sulfides. The oxide ore has been partially stripped, and 25 million tons having a copper content greater than 0.6% have been stockpiled. The stripping of oxide ore to expose sulfides will continue for several years. The copper oxide studies launched by Banner Mining in the 1950s were continued by Anaconda. Process routes investigated included: acid leaching; cyanide leaching; segregation roasting; chlorination, oxidation, and reduction; flotation in combination with leaching; cal¬cite flotation plus leaching; copper oxide flotation; and sulfidization flotation. By late 1968 test work favored a low temperature roast followed by an ammonia-ammonium carbonate leach of the ore. During this period, acid leaching had been discounted because of excessive con¬sumption of expensive acid by the high limestone ore matrix. Restric¬tive SO2 emission controls in 1971 suggested an increase in acid production and availability at a lower cost. Research then switched to acid leaching, resulting in development of the present flowsheet. Mechanically agitated leaching in a cascaded series of tanks was selected in preference to vat leaching after piloting a stirred-tank system. Preliminary testing of vat leaching indicated that inter-vat solution clarification problems could be caused by gas evolution from leaching of high limestone ore. Lack of research time for resolution of this problem, plus success of the pilot agitation system, dictated the final flowsheet selection. In 1973 Anamax decided to proceed with agitation acid leaching to produce liquor for a solvent ion exchange (SIX) and electrowinning (EW) installation. At that time, such systems were proving their eco¬nomic viability, and they allowed direct recovery of high quality cath odes while bypassing the smelting and refining of lower grade cement copper precipitates. The Anamax oxide plant processes a grade of about 1% recovera¬ble copper from stockpiled sources and new mine production. Unit operations include size reduction, agitation leaching, countercurrent decantation washing of residue, pH adjustment of solution, clarifica¬tion, solvent ion exchange, and electrowinning of copper, with a mini¬mum of environmental impact. The feed to the primary crushers is composed of: 1) -10-in. stockpiled oxide ore, which has passed through the mine crusher and has been grade-segregated to allow blending for a consistent feed material. 2) -4-ft ROM oxide ore stockpiled in the early mining operation. 3) -10-in. ore currently being mined. Ore is delivered in 100-ton mine trucks to a dump pocket, where an apron feeder transfers the ore to an inclined 6-in. fixed grizzly. The grizzly oversize passes through a 48 x 60-in. primary jaw crusher, where it is reduced to -6 in. The fixed grizzly undersize and the crusher product are conveyed to a 4-in. vibrating grizzly. Oversize from this grizzly reports to a 7-ft Standard cone crusher, where it is reduced to -1 in. The crusher product and the vibrating grizzly undersize are conveyed to two 8 x 20-ft double-deck vibrating screens that operate in closed circuit with two 7-ft short head cone crushers. Screen undersize (-1/2 in.) is delivered to a covered conical fine ore stockpile with a designed live capacity of 15,000 st. Beneath the stockpile are ten Pioche belt feeders feeding two conveyor belts that supply the milling circuits. Each group of five feeders has two variable speed and three fixed speed drives. Grinding to 95% -48 mesh is done in two parallel lines, each consisting of an 111/1 x 18S -ft rod mill and a 12' x 30-ft ball mill in open circuit. Water is added to maintain a pulp density of 72% solids in the mills. Further water additions are made to the ball mill discharge sumps to create a 60% solids slurry for transfer to the leaching reactors. The nominal grinding rate is 440 tph solids. A splitter box on the product from the mills separates the incoming slurry into two streams, one for the leaching reactors and the other for the pH adjustment reactors. This split is determined by the lime content of the feed material and the pH of solution entering the pH adjustment reactors. Five 30-ft diam by 31-ft high mechanically agitated, rubber-lined leach tanks are arranged in cascade, with gravity transfer from one to another by enclosed launders. Concentrated sulfuric acid (93.2%) from storage is added to the pulp principally at the first leach tank, and the reacted slurry leaving the cascade after a 5-hr residence time at 50% solids is pumped to countercurrent decantation (CCD) wash thickeners.' At this point, most of the soluble copper has been leached from the ore. Up to 250 lb of sulfuric acid is consumed per ton of ore, the acid being supplied in tank cars from a nearby smelter. Leached slurry at 50% solids is washed countercurrently with return raffinate from the SIX plant in a series of four Doff-Oliver center caisson, 400-ft diam thickeners. Solids advance from thickeners No. I through No. 4 to the tailings dam. SIX raffinate advances from thickener No. 4 to thickener No. l, where a pregnant solution overflows to pH adjustment. To optimize solvent extraction, the pH of the pregnant solution from CCD thickener No. I is adjusted from 1.5 to 2.5 by adding unreacted ground slurry (split from the grinding circuit) to the preg¬nant overflow of No. I CCD thickener. The resulting slurry is retained for 45 min in three mechanically agitated leach tanks. CO, generated during leach and pH adjustment is ducted to a demister, where entrained acid droplets are removed. The cleaned CO, and the No. 4 CCD underflow are the only effluents from the plant. The underflow of No. 4 thickener mixes with tailings from the sulfide concentrator and is neutralized. Slurry exits the pH adjustment vessels at about 10% solids and

Jan 1, 1985

By John C. Machisak

THE INJURY data and experience at quarries and related plants of the United States are presented in this publication under the following general sections: (A) General injury experience: Tables 1 through 3: Contain current data on the number of injuries, injury-frequency rates, location, accident agency, degree of injury, and related employment data. (B) Injury experience at different kinds of quarries: Tables 4 and 5: Contain current, data on number of injuries, injury-frequency rates, locution, accident agency, degree of injury, and related employment data. (C) Injury experience by States: Tables 6 through 19: Contain current data on the number of injuries, injury-frequency rates, accident agency, degree of injury, and related employment data.

Jan 1, 1958