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By Thomas W. Thompson

The Zebulon manganese deposit has been known for a number of years. It is situated about 3 miles east of Zebulon, Wake County, N. C., on the Avon, Privett property (fig. 1). The deposit outcrops as a vein, or dike averaging-16 to 20 feet in width. Drilling by the Bureau of Mines has shown that this outcrop results from consolidation of a number of small stringers near the surface. The surface width and the Percentage of manganese oxides (resulting from. weathering) diminish with depth. Below 50 to 80 feet, the vein material is mainly quartz and spessartite garnet. The deposit was first examined by W. T. Benson, a Bureau of Mines engineer, in 1941. The Bureau conducted a diamond-drilling exploratory program an the property from November 7, 1944, to January 25, 1945. A. total of 2,508.5 feet was drilled in 14 holes along the strike of the deposit.

Jan 1, 1950

By LeRoy D. Loy

To facilitate effective land reclamation and comply with existing state regulations, surface coal mine operators have in recent years adopted new mining techniques and equipment utilization for steep-terrain mining. In contrast to direct spoiling with draglines or shovels, these new mining methods require some form of overburden haulage from the excavation face to the point of reclamation. Although current mining methodologies such as Block-Cut, Modified Block-Cut, and Mountaintop Removal with Head-of-Hollow Fills have proven effective in achieving reclamation objectives, some production problems have resulted from pit congestion and excessive equipment delays. Recognizing this problem, the United States Bureau of Mines contracted Skelly and Loy to design innovative mining concepts utilizing conveyor haulage in place of trucks, loaders, dozers and scrapers. Project objectives required definition of current bottlenecks and development of mining systems to: alleviate bottlenecks; increase production potential; facilitate additional resource recovery; enhance reclamation; reduce safety hazards; and improve mining economics. Successful completion of project goals within a four year program involves systematic implementation of the following separate work phases: 1) Develop Concepts for Block-Cut Mining Systems; 2) Determine the Economic and Technical Feasibility of Concepts; 3) Prepare Detailed Design and Specifications of Concepts; and 4) Prepare Specifications for Conducting Field Tests of the Most Promising Concepts. At the time of this publication, Phases 1 and 2 are finalized with Phase 3 efforts well underway.

Jan 1, 1977

By M Ind, B Usher

Environmental regulators are increasingly adopting risk-based approaches to protect aquatic ecosystems. The Australian and New Zealand Environment and Conservation Council (ANZECC) and the Agriculture and Resource Management Council of Australia and New Zealand (ARMCANZ) (2000) introduced a risk-based approach to water quality guidelines that use decision frameworks to adapt water quality guidelines to local environmental conditions. Environmental quality objectives are no longer a simple ‘pass or fail’ number such as restricting total suspended solids (TSS) concentration to a specified limit for a particular rainfall event. This approach ignores spatial and temporal variability and the complexity of aquatic ecosystems. The risk-based approach establishes levels where adverse effects may occur. If these objectives are not met, further risk-based investigation is undertaken to assess the potential impact to the ecosystem. Klohn Crippen Berger (KCB) developed a dynamic simulation to predict the variability of sediment discharges, which is applicable for risk-based impact assessment of aquatic ecosystems. The application of the methodology over the life of a mine site is illustrated using a proposed South Pacific mine.The site location is in extremely rugged terrain with a high annual rainfall of approximately 4.7 m. There were two major sediment ponds proposed to reduce sediment discharges. A sediment and hydrology model was developed in Goldsim to provide daily estimates of run-off flows and sediment concentrations. A modified algorithm based on the Australian water balance model was created and incorporated into the Goldsim model and calibrated using recorded field data.The sediment module was developed based on a Revised Universal Soil Loss Equation (RUSLE) assessment, an annual prediction of soil loss over a catchment. A land-use type erosion vector was developed for different land-use types and for each of the subcatchments, a subcatchment erodibility vector was calculated. This vector and the daily rainfall erosivity factor were used to estimate the daily soil loss. The estimated soil loss and the modelled run-off yielded predictions of daily sediment concentrations. Finally, a sediment settling module was used to predict the daily efficiency of the sediment ponds and hence discharged sediment concentrations.The results of this modelling form the first part of a quantitative risk-based assessment into suspended solids impacts. The second part is the assessment of the effects of these predicted TSS concentrations on aquatic ecology. As part of the feedback-loop required in a risk management framework, if the environmental quality objectives are not met then further management decisions and actions may be required to reduce the risks presented by suspended solids. This may include modifications to: the mine-development plan; construction techniques and scheduling; processing techniques; rehabilitation strategies; or hard-engineering solutions such as sediment basins or flocculation.The Goldsim water-balance model provides a quick and efficient method of testing potential remedial actions on predicted sediment loads and concentrations. In combination with a cost-benefit assessment, options for reducing sediment releases from the mine site can be effectively and quantitatively compared thereby providing mine operators with a basis for decision-making. Regulators can also be satisfied that potential impacts to aquatic ecology have been comprehensively assessed and the necessary management decisions have been initiated to limit risk to the environment.CITATION:Ind, M and Usher, B, ?2014. A dynamic sediment simulation to predict the variability of suspended solids concentrations over life-of-mine for application to a risk-based environmental assessment, in Proceedings Life-of-Mine 2014 , pp 191–204 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By N. Hosking, M. Nichols

Aluminum alloys are increasingly used for automotive body construction but are susceptible to filiform corrosion or paint blistering in service. Filiform corrosion resistance of two 6000-series aluminum alloys was investigated in this work. Test panels were produced with full automotive coating layers and scribed parallel and perpendicular to the rolling direction. The corrosion exposure comprised an inoculation step with acidified salt solution followed by a 6-week exposure period at 95% relative humidity and either 25°C or 70°C. Greatest corrosion activity was measured on the alloy with higher copper content, AA6111. Higher temperature exposure led to greatly enhanced corrosion activity for AA6111 compared to the lower temperature exposure. The 70°C exposure also led to greater mean corrosion-affected area measurements for the low copper aluminum alloy, AA6022, but the effect was not as stark as that observed for AA6111. Rolling direction affected the propagation direction, and possibly mechanism, of the corrosion effects. Cross-sectional analysis revealed dense corrosion products within the filament, comprising mostly aluminum and oxygen but with traces of chloride and sulphate. Development of varying filament and blister types on a single test panel and the progression of corrosion underneath a filament into the substrate indicate that several corrosion mechanisms can develop on the test panel, testament to the complexity of the corroding surface developed on an aluminum automotive body panel.

Jan 1, 2015

By LESTER W. STRAUSS

OUR early knowledge of history and geography attracted most of us to the mineral resources of South America. The romantic tales of the Spanish activities, which were curiously alluring, and Prescott's exceptional bi-Conquest histories-Mexico and Peru-no doubt fired our minds with the .then metallic wealth of the Spanish colonies. Later developments, intensified within the past eighty years, have, however, changed its fame which now rests on such metallics as copper and tin, and among the non-metallics on nitrate of soda and petroleum. South America proper, excluding Panama, is made up of thirteen countries, a coincident parity with our original thirteen States. Its area is given as 7,231,300 square miles, or slightly less than North America (8,037,000 square miles), is made up as follows:

Jan 1, 1930

By Arthur L. Walker

DURING my trip around the world last year, covering a total of 45,000 miles, I saw many things of especial interest from an engineering viewpoint. Sailing from New York, I went through the Panama Canal, down along the west coast of South America, then over the Peruvian and Bolivian Uplands; thence to the celebrated copper mines of Chile; across the Andes, to Buenos Aires, then to Rio Janeiro; across the Atlantic to Lisbon, Portugal; through Spain to the Rio Tinto mines; across southern Europe to Italy; thence to Alexandria and Cairo; up the Nile to Assouan, to Palestine through the Suez Canal, Red Sea and across to Bombay; up to the country of the Moghul Emperors; then to Calcutta, from which point I shipped to Singapore, as having crossed the Equator twice, I wished to shake hands with it again; thence to Shanghai and Pekin; up to the Great Wall of China, back to Nanking and up the Yangtse to Hankow; then back and eastward across the Pacific Ocean. The climatic changes were sudden and severe; there was winter where, for me, it should have been summer and vice versa. The weather man was very kind, however, and until I was near the Aleutian Islands there were only three rainy days in a total of 220.

Jan 1, 1924

By G. Jandieri

An improved methodology is presented for assessing the economic feasibility and effectiveness of recycling industrial waste. The methodology is based on the break-even control mechanism, but at the same time provides for the introduction of new evaluation criteria such as the threshold of conditionality and the degree of ore substitution. Based on an improved analysis, it becomes possible to more precisely predict recycling efficiency. A more refined determination of the lower limit of concentration of recoverable metals, at which technogenic waste can be assigned the status of secondary raw materials and processed profitably, leads to a significant expansion of the secondary raw material base suitable for recycling. The potential for recycling manganese-containing dust from the production of ferrosilicomanganese, dehydrated sludge from the hydro separation of slags, and cake from the production of electrolytic manganese dioxide at the Chiatura mining enterprise is used as an example. It is shown that with a threshold of –24% Mn content, the highest recycling efficiency can be achieved by the production of low-phosphorus manganese slag and conversion to ferrosilicomanganese using the above waste to replace 40-60% grade III and IV manganese concentrates in the feed.

Jan 26, 2023

By J. E. Udd, S. Bharti

"ABSTRACTMonitoring of stability conditions in underground mines permits management to assess the response of the rock mass to mining strategies. Monitoring of movements, if sufficiently early, may also permit operators either to avert potential failures or to develop alternative approaches to extraction.The Falconbridge organization has been a leader both in the uses of rock mass stability monitoring techniques and in the development of instrumentation.In this paper, the authors review the Falconbridge approaches to monitoring, and present case history examples to illustrate the application of the various instruments used.RESUMELa surveillance des conditions de stabilité dans les mines souter-raines permet aux chefs d'enterprises d'évaluer les réactions de la masse rocheuse aux méthodes d'extraction minière. Si le contrôle des mouvements du terrain est effectué assez tot, l'exploitant pourra prendre les mesures qui s'imposent, soit pour éviter les détériorations possibles ou pour mettre au point de nouvelles méthodes d'extraction.Le groupe de la Falconbridge a ére le premier a utiliser des techniques de vérification des conditions de stabilité de la masse rocheuse et a mettre au point l'appareillage requis.Ce document présente une revue des méthodes employées par la Falconbridge pour assurer la surveillance active de la stabilité de la masse rocheuse et illustre au moyen de cas concrets la manière d'utiliser les divers appareils."

Jan 1, 1987

By James Douglas

IN discussing the waste upon which hinges, or is supposed to hinge, so largely the preservation of our national resources, the conclusions reached would be more reliable if actual ex¬perience were consulted, and fewer deductions were drawn from general statements, which are often the product of the imagination. It cannot be questioned that the value of by-products has not been sufficiently appreciated by us, and that our, tardiness in recovering the useful ingredients of the escaping gas of our coke-ovens is one of the most glaring instances of shortcoming in that direction. And yet even for that sin there is some palliation in the immature condition of affiliated industries. I presume that it is admitted, without argument, that, except under very exceptional conditions, all the elements cannot be recovered from most of the ores or natural products. which we treat. While it is a shame that the by-products from our coke-ovens should be dissipated, Edward W. Parker's report to the U. S. Geological Survey for 1906' supplies a fairly good excuse in justification of this appalling waste. He says (pp. 773 to 774.)

May 1, 1909

By A E. A Northcote

The drive to maximise performance, reduce maintenance times, improve safety of operation and maintenance on mining equipment while keeping both operational and capital costs down has provided good nourishment to the creative thinking of designers of skip loading facilities for hoisting systems. Early systems were as simple as manual loading which evolved into a surge bin feeding a skip through a chute finally maturing into a measuring flask discharging into a chute to feed the skip and more recently high capacity skip feed conveyors. This paper will compare the concepts of the skip loading flask and the skip loading conveyor systems over a range of winding distances and hoisted tonnages. The comparison will draw on actual installations and review operational, maintenance and safety. An order of magnitude capital cost estimate will be used to complete the financial comparison. This will enable the reader to understand the concepts and costs and assist in selecting the appropriate skip loading system.

Mar 21, 2011

By Charles A. Kumke, GROVER J. DUFF

This paper describes mining and milling practices at the San Xavier mine of the Eagle- Picher Mining & Smelting Co. in Pima County , Ariz . It is one of a series being prepared by the Bureau of Mines on mining methods in various mining districts of the United States . The mine and the mine surface buildings are situated in sec . 11 , T. 17 S. , R. 12 E. , in the Sierrita Mountains in Pima County , 20 miles south by southwest from Tucson via the San Xavier Mission and Twin Buttes road ( fig . 1). A well-maintained dirt road continues southward toward Twin Buttes for 4 miles and then branches off northeasterly for 5-1/2 miles to Sahuarita , where the concentrator of the Eagle - Picher Mining & Smelting Co. is situated on the west slope of the Santa Cruz River Valley . Tho mill is connected by a railroad branch to the main line of the Southorn Pacific Railroad between Tucson and Nogales . It is generally assumed that the San Xavier silver - lead deposits became known to the Jesuits and Spaniards in the early part of the eighteenth century . However , no extensive mining operations were conducted by the Jesuits in what is now Arizona before the middle of the eighteenth century . The completion of the Southern Pacific Railroad from Los Angeles to El Paso in 1881 marked the beginning of intense activity in mining throughout the southern part of the State . The San Xavier mine produced 6,000,000 pounds of lead , considerable zinc ( exact figures not available ) and $200,000 in silver from 1897 to 1918 . The Eagle- Picher Mining & Smolting Co. re- opened the San Xavier mine in 1943 and built a 175-ton concentrator at Sahuarita . The capacity of the plant was increased in 1944 to 500 tons in order to take in custom ore . The surface rocks at the San Xavier mine are , in great part , of sedimentary origin . Limestones of various ages comprise most of the exposed sedimentaries , some quartzites and shales also being present . The sedimentaries rest upon coarse - textured , intrusive , igneous rocks comprised mainly of granite and granodiorite . The San Xavier deposit is composed of galena , sphalerite , chalcopyrite , and pyrite . Generally the individual ore bodies are small , discontinuous , and occur as irregularly spaced shoots or lenses . The present mine workings extend horizontally a distance of 2,700 feet and vertically 900 feet . Owing to ground weight, all mine openings are timbered on the levels and all stopes are square- setted and waste -filled .

Sep 1, 1950

By J D. Grobler

In the heavy minerals industry wet high intensity magnetic separators (WHIMS) are widely applied for primary magnetic separation of magnetic minerals and non-magnetic minerals. This piece of equipment is well known for its robustness to handle a wide range of feed conditions and still produce stable product qualities and yields. A heavy mineral production plant aimed to increase production through their WHIMS plant containing 11 duties and 26 units. Instead of adding new units an alternative approach was followed to firstly maximise the efficiency of the current circuit and secondly to consider possible reconfiguration of the circuit to produce additional product tons for minimal capital investment. In order to characterise the current circuit’s efficiency, identify pockets of excellence and determine the areas for improvement the whole circuit was surveyed several times and mineral mass balances were drawn up.This paper aims to: discuss the modelling process of the WHIMS circuit,explain the calibration and validation process followed to ensure the accuracy, anddemonstrate the sensitivity analysis performed on the different options. From this analysis an efficiency criteria was developed to compare different circuit configurations that enabled the production plant personnel to make an informed decision on the way forward.

Oct 5, 2011

By P Jones

Fosterville Gold Mine is an underground operation located approximately 20 km northeast of Bendigo and 170 km north of Melbourne. Fosterville is the most productive gold mine in Victoria, producing over 100 000 ounces of gold annually. Access to the underground is via two portals cut into previously mined open pits. Fosterville is 100 per cent owned and operated by Northgate Minerals, a gold and copper producer with mining operations, development projects and exploration properties in Canada and Australia. (Northgate Minerals Corporation, 2010).Underground operations commenced in 2006 with ore production focused on three orebodies – Phoenix, Ellesmere and Falcon. Primary ventilation to the existing operation is provided by wall mounted exhaust fans located underground. This system provides approximately 220 m3/s of primary ventilation to the operations with intake achieved through the portals. Exploration drilling during recent times has uncovered another lode, the Harrier orebody, located 2.2 km to the south of the existing workings. Due to the location of the new Harrier orebody, vent analysis revealed the requirement for a separate source of primary ventilation to service Harrier. Figure 1 shows a current underground long section and the location of Harrier in relation to the existing workings. A number of options were considered, including a parallel decline from one of the existing two portals, however, a 4 m diameter, 350 m deep surface ventilation shaft was selected as the most suitable and feasible option for the mine.In December 2009, geotechnical drilling commenced to determine the most suitable location for the shaft. Site locations were constrained by a number of factors including interaction with current and planned surface infrastructure as well as proximity to nearby community members at the southern end of the mining lease. Results from the geotechnical drilling indicated that the majority of the proposed shaft would be suitable for a 4 m raise bored shaft. However, the top 40 m of the proposed shaft were deemed unsuitable for raiseboring due to zones of weakness that would likely influence the stability of a 4 m diameter shaft (Robinson, 2010). Following independent geotechnical advice and input from experienced raiseboring contractors, it was decided to construct a 40 m presink prior to raiseboring activities.A crane mounted auger was chosen as the preferred method to construct the presink. This method differs substantially from traditional blind sinking methods which are highly labour intensive. The auger, or drill, is designed and built to be attached to a crane base using engineered pivoting cradles. The auger is suspended from the crane jib by pennant ropes attached to hydraulic rams. The auger applies rotation and axial loads to the drilling tools via a telescopic steel bar (Kelly Bar). Drill tools are attached to the bottom of the Kelly Bar and comprise an assortment of drilling buckets and reaming barrels. Drilling buckets cut in situ material and collect the spoil for removal and are used for the excavation of the ‘pilot hole’. Reaming barrels are used to expand the diameter of the excavation by directing cuttings into the pilot hole which are then removed by the drilling bucket. In June 2010, the presink was constructed utilising a crane mounted auger with support to the shaft walls applied via a remote fibrecrete unit. The excavation method involved cutting a series of smaller diameter holes (2.4 m wide) before reaming out to the final diameter (4.7 m wide) in 4 - 5 m vertical increments. Fibrecreting activities were carried out on the backshifts, to ensure the stability of the excavation and minimise interaction between the two units. The presink was completed on time and offered many advantages compared to traditional blind sinking methods, including safety, costs and timing. This paper will examine the benefits of this method as well as lessons learnt and future applications in the mining industry.

Mar 21, 2011

By T. Mapeta, S. M. Rupprecht

Discussions on reasonable prospects for eventual economic extraction (RPEEE) include the CRIRSCO 2019 reporting template, JORC 2012, SAMREC 2016, CIM 2019 Guideline, and Regulation SK-1300. These reporting codes all conform to the principles of reasonable prospects for eventual economic extraction. To demonstrate that all mineral resources align with the RPEEE principles, the Competent Person is responsible for endorsing the mineral resource estimate. Moreover, establishing mineral resources and RPEEE requires technical input from numerous technical specialists, such as mining and geotechnical engineers, metallurgical engineers, social and environmental specialists, financial and governmental practitioners, and more. The aim is to improve the transparency of the disclosure of mineral resources and guarantee the investor or potential investor that the reporting of mineral resources is based on reliable technical, economic, environmental, social, and other relevant factors. This paper discusses the important protocols and definitions in some international reporting codes for the purposes of developing a guide memoire that will assist the Competent Person to interpret, establish and apply RPEEE appropriately and easily.

Jun 27, 2025

By S. C. Lindsay, D. J. I. Evans

With each year that passes hydrometallurgical processes are being more widely used to recover base metals from ores and concentrates. Generally these processes involve liquid-solid separation of metal-bearing liquors from barren residues. This may be done by countercurrent decantation, thickening and filtration, or filtration or centrifuging only. Choice of method is governed by chemical and physical characteristics of the pulps; cost of labor, power, materials, and waste disposal; and even by the amount of capital available when building plans are under way. Since all these controlling factors change with time and location and with technological advances, the separation method must be re-evaluated when new plants or expansion programs are contemplated. At the Sherritt Gordon refinery in Fort Saskatchewan, Alberta, an ammonia pressure leaching process recovers nickel, copper, and cobalt from the nickel concentrate produced at Lynn Lake, Manitoba.' Pregnant liquors are separated from the barren leach residues in a thickener-filter liquid-solids circuit. The decision to use this type of separation was based on a summation of economic factors during the pilot plant and design stages. Although countercurrent decantation was not incorporated in the refinery flowsheet, the problems solved in thickening the leach residues and the experience gained in operating the commercial plant are pertinent to countercurrent decantation, particularly where it is applied to slow-settling leach slurries or slurries having a relatively high vapor pressure of a valuable, obnoxious, or poisonous vapor. The nickel sulfide concentrate is leached under pressure with air and ammonia to dissolve the nickel, copper, and cobalt and most of the sulfur, leaving a residue of iron and siliceous material to be discarded.2,3 During a subsequent distillation stage to recirculate part of the ammonia used for leaching, copper is recovered from the solution as a high-grade sulfide. After further purification to remove traces of copper and undesirable sulfur compounds, the nickel is precipitated by hydrogen under pressure at an elevated temperature. The cobalt, together with unreduced nickel, is precipitated as a mixed metal sulfide by hydrogen sulfide for eventual cobalt recovery. The end solution is then evaporated to produce a crystalline ammonium sulfate byproduct. As shown in Fig. 1, the nickel concentrate is leached in a continuous, two-stage, countercurrent circuit. In the first stage, or adjustment leach, the fresh concentrate is partially leached (60 pct) in the liquor produced in the second stage, or final leach. This system is necessary to produce a solution suitable for treatment in the copper sulfide precipitation step. Both stages are carried out at 170" to 180°F under a total pressure, with air, of 100 to 120 psig in large, horizontal, water-cooled autoclaves. The autoclaves are divided by overflow weirs into four compartments to give a certain amount of staging during leaching. In each compartment a relatively slow-speed (100 to 150 rpm) agitator maintains the nickel concentrate in suspension. Adjustment leaching, carried out in two autoclaves in parallel, is followed by a thickener-filter liquid-solids separation. Final leaching is done in two parallel trains of three autoclaves using aqueous ammonia to extract metal values from the adjustment leach residues. This too is followed by a thickener-filter circuit. Here the residues are repulped once with aqueous ammonia and twice with process water and refiltered after each repulping stage, to insure a maximum recovery of soluble metal, before being discarded to the tailings pond. The adjustment leach is followed by a seven-disk Dorr-Oliver American filter, with an agitator in the filter boot to keep the residues in suspension. Agitators have also been installed beneath the four ten-disk filters following the final leach thickener. Slurry Characteristics: A typical analysis of the Lynn Lake nickel concentrate, produced by differential flotation,' is given in Table I. The concentrate comprises 35 to 49 pct pentlandite, (Fe, Ni)S; 3 to 6 pct chalcopyrite, CuFeS,; 1 to 4 pct pyrite, FeS2; and 24 to 42 pct pyrrhotite, Fe1-2S (where x varies between 0 and 0.2) together with various siliceous minerals such as actinolite, hornblende, and talc. During leaching, the sulfur in the pentlandite, chalcopyrite, and part of the pyrrhotite is successively oxidized through the thiosulfate and poly-thionate radicals to sulfate and sulfamate." Simultaneously the nickel, copper, and cobalt are converted to soluble ammine complexes and the iron in the pentlandite, chalcopyrite, and part of the pyrrhotite is oxidized to a spongy hydrated iron oxide. The hydrated iron oxide is readily broken down by the agitation needed to sustain the concentrate in suspension and to maintain a satisfactory mass transfer of the oxygen from the air to the slurry. Table I shows typical chemical, screen, and

Jan 1, 1961

The domestic mineral and material industry grew in some key sectors and declined in others during 1993; overall, the industry was poised at yearend for stronger growth in 1994. Metals production is expected to increase with anticipated growth in U.S. motor vehicle sales. There also may be increased demand for industrial minerals in 1994 based on domestic highway and building construction. Additional information on events, trends, and issues in the mineral and material industry is presented below and in the commodity sections that follow. Overall Industry Performance The value of processed materials produced from minerals in the United States during 1993 is estimated to have been about $326 billion, an increase of 5% compared with that of 1992. The estimated value of U.S. raw nonfuel minerals production in 1993 was $31.6 billion, about 1 % less than in the preceding year. The value of domestic minerals production has risen in 28 of the last 33 years.

Jan 1, 1994

1. Domestic Production and Use: In 1974 twelve companies operated 31 primary aluminum reduction plants, three firms accounting for 64% of production. Washington, Oregon, and Montana accounted for 32% of the total production; Gulf Coast States, 24%; Northern States, 33%, and other Southern States 11%. Output of primary metal in 1974 was valued at $3.4 billion. About 80% of the secondary recovery was from new aluminum base scrap and 20% was from old scrap. Aluminum consumption by 25,000 firms was centered in the East Central United States. Of aluminum consuming industries in 1974, building accounted for about 27%; transportation 19%; electrical, 13%; packaging, 15%; and consumer durab1es, 9%.

Jan 1, 1975

By Forrest T. Moyer

The overall injury experience, excluding officeworkers, at metal mines and mills during the period 1961-63, reported by operators, was 174 fatal and 11,357 nonfatal disabling work injuries during an aggregate worktime of 451 million man-hours. These injuries occurred at a frequency rate of 25.55 per million man-hours, comprising 0.39 for the fatal and 25.17 for the nonfatal injuries. The combined severity rate of the disabilities was 3,779 days lost or charged per million man- hours, consisting of 2,314 for the fatal and 1,465 for the nonfatal injuries. The average severity of all disabilities during 1961-63, was 148 days per injury. The number of days lost and/or charged, ex- cluding officeworkers, at metal mines and mills during the period 1961-63, was 1,440,000 days charged to fatal injuries and 661,290 days lost from nonfatal injuries. An average of 70,966 men were at work daily. The outstanding cause for fatal injuries was the fall of rock or ore from the back, while most of the nonfatal injuries were caused by handling materials, slips or falls of persons, machinery, falls of rock or ore from back and side, and haulage. Injury experience at primary nonferrous smelters, refineries, and reduction plants during 1961-63, excluding officeworkers, was 23 fatal and 3,214 nonfatal disabling injuries during a total work- time of 277 million man-hours. The injuries occurred at frequency rates of 0.08 per million man-hours for the fatalities and 11.61 for the nonfatal disabilities. The severity rate for all injuries of 1,085 days lost or charged per million man-hours comprised 498 for fatal and 587 for nonfatal injuries. The number of days lost, excluding officeworkers, at primary nonferrous smelters, refineries, and reduction plants during 1961-63 was 138,000 from fatal injuries and 162,385 from nonfatal injuries, An average of 34,938 men were at work daily an active plant days during the three year period. The injury experience for officeworkers at metal mines and mills during the period 1961-63 was 20 nonfatal disabling work injuries, (there were no fatal injuries reported), during an aggregate work-time of 28,694,244 man-hours. These injuries occurred at a frequency rate of 0.70 per million man-hours. There were no fatal injuries reported for officeworkers at non-ferrous smelters, refineries, and reduction plants during 1961-63, but seven nonfatal injuries were reported during an aggregate work-time of 30,814,772 man-hours. The frequency rate for these injuries was 0.23 per million man-hours for this period.

Jan 1, 1967

By Ricardo Oliveira Gallart de Menezes, Claudio Gerheim Porto, Reiner Neumann

This study presents the results of the mineralogical characterization of Ti-Zr ore from the Jiboia deposit, located in the municipality of Brejo Grande, Sergipe State, Brazil. The goal is to identify the main Ti, Zr, and REE carrier phases in the different particle size fractions in order to subsidize the development of a cost-effective processing route for the production of heavy minerals concentrate. The mineral resources are estimated at 31.5Mt @ 3.5% of total heavy minerals (THM) for the Jiboia Ti-Zr deposit, of which 5 Mt @ 3.39% THM were classified as measured and indicated. The mineralogy of the heavy mineral concentrate recalculated to 100% is represented by 22% pseudorutile, 20% ilmenite, 15% zircon, 5% rutile, 2% monazite, and 36% gangue. The ore minerals are preferably concentrated in the 150 × 53-μm granulometric fractions, which represents 80% of the THM and only 27% of the ROM (Run-of-Mine) ore. Therefore, by cutting the ore to 150 μm, it is possible to discard 73% of the ROM mass and recover 80% of the THM, including 100% of the zircon and monazite which are the most valuable minerals present in the heavy mineral concentrate. The global recovery of THM starts to decline rapidly in the fractions finer than 106 μm and the mass disposal (tailing) does not follow in the same proportion. Higher zircon contents in the concentrates of the fractions finer than 150 μm imply a significantly lower recovery of heavy minerals like ilmenite, pseudorutile, and rutile that make the fractions coarser than 150 μm.

Jan 9, 2024

Fresh coal refuse was sampled at 21 US mines. Five of the sites were drilled, sampled and tested more extensively. Field tests included Standard Penetration tests, and in-situ density determinations. Laboratory testing included specific gravity, gradation, maximum, and minimum density tests, as well as 9-inch) drained triaxial tests, 9 and 10-inch, permeability tests, slake-durability tests, and accelerated slaking tests. Some refuse sites had significantly lower shear strengths and permeabilities in old, slaked or weathered material than in fresh refuse. Testing did not show useful correlations between strength and permeability changes and other material properties which could serve as "predictors". Testing of refuse under "accelerated slaking" conditions showed promise as a comparative tool and as a potential means of preparing artificially aged samples for evaluations of long-term behavior. The state-of-the-art in slaking research, and the implications of slaking on coal refuse disposal, and on practical methods of managing slaking, were also studied.

Jan 1, 1981