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By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines or phosphate ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pct. When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900" to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By F J. van Zyl, C A. S Thomson

"Anglo American Thermal Coal conducted a high level study of all of its underground collieriesto determine if there were electrical power saving opportunities at the main fan stations.A systems approach to electrical power savings was used to identify potential savings in the various components of a primary ventilation system. Furthermore, only technological interventions were considered to obtain electrical power savings. From generic technological driven electrical power saving strategies four energy savings strategies were identifi ed, namely potential overall reduction in ventilation volume, improved underground air utilisation, optimisation of fan-set effi ciency and potential fan drift design optimisation. The study indicated the colliery (say Colliery D, Shaft A) that showed most potential in the four technology driven energy savings strategies. A detail study was undertaken at Colliery D, Shaft A to increase the confi dence level and the study confi rmed that up to 45 per cent of the main fan absorbed power can be saved (475 kW) with an estimated capital payback period of 18 months. The study clearly showed that an integrated approach to optimise the entire mine ventilation system has great potential for power saving, and that a clear, well -defi ned methodology will be applicable to similar underground collieries. This paper discusses the background to the study and the process followed to implement the energy savings initiative at Colliery D, Shaft A."

Jun 1, 2012

By Carl Zapffe

JUST as 85 per cent of the total ore produced annually in the United States comes from the Lake Superior region, so does one of its six producing districts-the Mesabi --dominate that region both as to production and total reserves; also in ore Treatment, Mesabi practice is dominant. Therefore, in discussing future reserves or ore treatment one may visualize one big district with five smaller districts as so many satellites. In considering the future less thought need be given to processes than to, economic policies. For thirty years experimental work has now been in progress and every possible ore-treatment process now known has been given attention. This work was started when there was no recognizable danger of running short of ore because of rapid consumption. Thirty years ago the annual production had just progressed beyond the twenty-million bracket for the first time. Today fifty million tons is considered a small production, and last year the aim was for 65 million. At the recent rate of production the life of the estimated reserves of good ore is about 35 years. All who know the local terrain are convinced that, unlike thirty years ago, the chances for discoveries of new big deposits are nil; treatment of lean ores obviously is to become an all-important topic.

Jan 1, 1938

By O. A. Olaseinde

The 2101 lean duplex stainless steel has wide potential application in storage, heat exchangers, and in the oil and gas industries. This work investigates the electrochemical behaviour of 2101 duplex stainless steel with an addition of 0.15 wt% ruthenium, using potentiodynamic techniques in 1M H2SO4 at 25º, 40º, 60º, and 80ºC. The microstructures of samples were characterized using optical metallography and scanning electron microscopy. The results showed that the ruthenium addition moved the corrosion potential of alloy 2101 to a more positive potential. All samples containing ruthenium displayed longer passive regions at 25°C and 80°C compared to those without ruthenium. Alloys without ruthenium had higher critical current density (icrit) values when compared to the alloys with ruthenium. Ruthenium additions decreased the passive current densities and inhibited anodic dissolution. At room temperature the corrosion rate of alloys with ruthenium was lower than these without ruthenium.

Jan 1, 2012

By C. H. Snyder

Tlie paper deals with details of construction of the Colmol, including improvements in design that will be incorporated in new models. These improvements are results of problems encountered and worked out in experimental operation with the unit. Also included is a descriptive plan of operating where roof and air conditions are of the worst. A summary of production results from the machine is presented. THE various units used in conventional mining are well-built. Generally the universal under -cutter, driven by approximately 50 hp, weighs approximately 25,000 lb, and is a very strong piece of machinery. The drilling machine is not as husky as an undercutter, generally is driven by approximately 15 hp, weighs about 9000 lb, but is a well-constructed, sturdy mechanism. The loader, usually 60 hp, weighs approximately 24,000 lb, and certainly is a strong, rugged unit. The working parts of all three of these sturdy conventional units have been made of tough steel alloys for some time. With the use of three machines having a combined horsepower of approximately 125, with a combined weight of approximately 60,000 lb, and with the possible inherent advantage of single purpose machines performing a single function, the use of considerable explosives has been required to complete the conventional mining cycle. It was therefore our thought, upon conception of the first Colmol,* a machine which was to do the work accomplished by the three conventional units plus explosives, that it must be a rugged mechanism. Now, after building several units, after going through many and varied experiments, and after the expense of some million-odd dollars, we find we have observed and learned nothing that contradicts our original thinking. A continuous miner, if it is to give satisfactory performance, must be a rugged piece of machinery. Ruggedness has been, and is constantly, in the minds of our engineers in designing the Colmol. Improvements where they have been required have been in the direction of more weight, more horse power, better alloys, bigger hydraulic pumps and motors; in short, always, toward more torque, greater strength, more ruggedness. Now not only the finest steel alloys are utilized but more than 100 Timken roller bearings are used. The following is some pertinent data on models being built and to be built: The low unit: Cutting height, 30 to 42 in. By changing the upper gear case the maximum cutting height can be increased to 61 in., giving a variable cutting height of from 30 to 61 in. Width of cut, 9 ft 6 in. Width of body, 75 in. over tracks. Height of body, 271/2 in. Overall length, 23 ft. Driven by three 50-hp motors or a total of 150 hp. Weight, 26 tons. The intermediate unit: Cutting height, 451/2 to 72 in. Width of cut, 9 ft 8 in. Width of body, 77 in. over tracks. Height of body, 43 in. Overall length, 26 ft. Driven by three 70-hp motors or a total of 210 hp. Weight, 35 tons.

Jan 1, 1951

By C. H. Snyder

Tlie paper deals with details of construction of the Colmol, including improvements in design that will be incorporated in new models. These improvements are results of problems encountered and worked out in experimental operation with the unit. Also included is a descriptive plan of operating where roof and air conditions are of the worst. A summary of production results from the machine is presented. THE various units used in conventional mining are well-built. Generally the universal under -cutter, driven by approximately 50 hp, weighs approximately 25,000 lb, and is a very strong piece of machinery. The drilling machine is not as husky as an undercutter, generally is driven by approximately 15 hp, weighs about 9000 lb, but is a well-constructed, sturdy mechanism. The loader, usually 60 hp, weighs approximately 24,000 lb, and certainly is a strong, rugged unit. The working parts of all three of these sturdy conventional units have been made of tough steel alloys for some time. With the use of three machines having a combined horsepower of approximately 125, with a combined weight of approximately 60,000 lb, and with the possible inherent advantage of single purpose machines performing a single function, the use of considerable explosives has been required to complete the conventional mining cycle. It was therefore our thought, upon conception of the first Colmol,* a machine which was to do the work accomplished by the three conventional units plus explosives, that it must be a rugged mechanism. Now, after building several units, after going through many and varied experiments, and after the expense of some million-odd dollars, we find we have observed and learned nothing that contradicts our original thinking. A continuous miner, if it is to give satisfactory performance, must be a rugged piece of machinery. Ruggedness has been, and is constantly, in the minds of our engineers in designing the Colmol. Improvements where they have been required have been in the direction of more weight, more horse power, better alloys, bigger hydraulic pumps and motors; in short, always, toward more torque, greater strength, more ruggedness. Now not only the finest steel alloys are utilized but more than 100 Timken roller bearings are used. The following is some pertinent data on models being built and to be built: The low unit: Cutting height, 30 to 42 in. By changing the upper gear case the maximum cutting height can be increased to 61 in., giving a variable cutting height of from 30 to 61 in. Width of cut, 9 ft 6 in. Width of body, 75 in. over tracks. Height of body, 271/2 in. Overall length, 23 ft. Driven by three 50-hp motors or a total of 150 hp. Weight, 26 tons. The intermediate unit: Cutting height, 451/2 to 72 in. Width of cut, 9 ft 8 in. Width of body, 77 in. over tracks. Height of body, 43 in. Overall length, 26 ft. Driven by three 70-hp motors or a total of 210 hp. Weight, 35 tons.

Jan 1, 1951

By Scott L. Burrill

In the past 100 years in the mining industry, many references are made to the "Widow Makers", 'Miners Con" and many other derogatory names related to physical hazards. They reflect the lack of understanding of the problem facing the early day miner. Workers were enticed, in some cases, to take a chance against a medical question mark that could not be explained by the doctors of those early days. The optimism of the mining man and the challenge of work led him to believe that the physical results could not apply to himself and only applied to some of his fellow workers. Much of the time, to his dismay, he found that he had lost to the game of chance. As medical knowledge improved and mining management began to understand the effects of conditions in the mines and beneficiation plants, steps were taken to alleviate or eliminate the health hazards that were present, The organizations engaged in mining that did not take an active interest in employee health found themselves to eventually be placed in the position of being forced to take action by Federal and State legislation. We find ourselves in the 1970's on the threshold of a new era where the general public is concerned vociferously about "their" environment and the industry feels the governmental controls more strongly than ever before. It can be said that in the past 100 years we have advanced from the "Widow Makers" of the 1870's to the "Retirement Makers" of today. The difficulties of the early days in underground mining were of times compounded in the beneficiation plants above ground. As late as the 1950's some crushing and milling installations in the United States were operating under environmental conditions that complied to the then existing legislation but were below the health level sought by conscientious management. The Mission Concentrator Crushing Plant was designed in 1959 and placed in operation in July, 1961. In the past ten years the crushing plant has been continuously complimented by technical people, government inspectors and casual

Jan 1, 1971

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines Or phosphate Ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pet' When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900° to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By Arshad Ali, Mudassar Khan, Waqas Ahmad, Nisar Muhammad, Muhammad Salman

Talc is an important industrial commodity due to its wide-ranging intrinsic properties; however, the presence of iron oxide hinders its industrial utilization. Herein, we report the purification of Pakistani talc through acid leaching assisted by chlorination techniques. Among various acids (i.e., HF, HCl, H2SO4, HNO3, H3PO4, Aqua Regia, H3BO3, and CH3COOH), HF was the most effectual, removing up to 67% of iron oxide at optimized conditions of temperature, 90–95 °C; time, 120 min; HF concentration of 5%; and at a pH of 1. In the second step, chlorination of the talc sample was performed separately by both wet chlorination and dry chlorination. Wet chlorination was carried out using KMnO4 and HCl in slurry phase which shown about 52% removal of iron oxide, whereas dry chlorination with CaCl2 at high temperature attained about 48% removal of iron oxide from talc sample. Further, the wet chlorination of HF leached samples was also performed under optimized conditions which showed up to 87% iron oxide removal. The original and treated talc samples were characterized by EDX and XRD analysis to monitor the changes in the elemental and mineralogical composition of the talc. It can be concluded that HF leaching followed by wet chlorination is an effective method for the upgradation of low-grade Pakistani talc ores.

Aug 9, 2021

By Konrad J. Banaszak

At least for the time being, these seems to be general but not universal agreement that Mississippi Valley type ore deposits formed entirely or in major proportion from heated connate brines (e.g. 1) Even with that agreement, major questions remain. I will present a coherent geochemical model and look at alternative views in the context of that model. The pattern of the following discussion is 1) the general geochemistry of the ore fluid, 2) the density-temperature relations and possible mechanisms of flow of those fluids, 3) the mechanisms of transportation and deposition of the ore minerals, and 4) a summary of a coherent model of genesis. The data equivocally define much about the deposits. The answer is still elusive.

Jan 1, 1979

By Jack E. Bailey

From a start in 1957, milling 3,000 TPD of copper ore, the Cyprus Pima Mining Company, which is located 20 miles south of Tucson, has expanded its operations four times, and is currently treating over 55,000 TPD of ore.1 The most recent expansion added a 14,000 TPD semi-autogenous concentrator to the operation. Shortly after startup in January, 1972, the semi -autogenous plant exceeded the designed tonnage rate.2 Ore from the mine is delivered to a 54 inch gyratory crusher and is reduced to a minus 8 inch product. From the crusher the ore is stacked on a surge pile and then reclaimed and fed to the two parallel grinding circuits. A 28 ft. diameter by 12 ft. long semi-autogenous mill and a 16.5 ft. diameter by 19 ft. long ball mill make up each grinding circuit. The semi-autogenous mills are in closed circuit with screens and the ball mill circuits are closed with hydro- cyclones. Rougher flotation follows the grinding operation. Concentrate regrinding and cleaner flotation is performed in Mill I, the older mill. The generalized flowsheet is shown in Figure 1.

Jan 1, 1975

By W. Wanjokow, H. O. Hofman

In the metallurgical treatment of most metallic sulphides it is usually necessary to carry on a roasting-operation. In some cases the raw ore will have to be roasted; in others the intermediary product, matte, produced in smelting the ore; again, both ore and matte may have to be subjected to oxidation at elevated temperatures. Most metallic sulphides, when given an oxidizing-roast, pass through the state of sulphate before they are converted into oxides. All metallic sulphates are completely decomposed by heat, with the exception of those of lead and bismuth, but the temperatures at which decompositions take place vary with the metals. Some metallic sulphates are converted directly into oxides, others first form basic salts. Alkali-earth sulphates show a behavior similar to that of metallic sulphates, while alkali sulphates do not part with their sulphur trioxide even at elevated roasting-temperatures. The older data on the decomposition of metallic sulphates by heat are only approximations. Thus Kerl' gives the order in which metallic sulphates are decomposed with increasing temperatures as (Ag2, Fe, Cu, Zn, Ni, Co, Mn, Pb) SO,. More recent statenlents give definite temperatures and modify the formerly accepted order of decomposition. But the temperatures given by various observers do not agree. The reasons for this are to be found mainly in the method and manner of carrying out experiments. Four methods have been used: 1. Heating in an evacuated tube; 2, heating in a tube closed at one end, without air-cur-

Jan 1, 1913

By G. R. Leverant, C. P. Sullivan

To evaluate the effect of a dispersion of nondeform-able, incoherent, second-phase particles on high-strain cyclic deformation and fracture, recrystallized TD-nickel (Ni-2ThO2) and a commercially pure nickel, Ni-200, were fatigued under strain control at total strain ranges varying from 0.009 to 0.036. Relative to the Ni-200, the slip at the surface of the TD-nickel was more wavy and discontinuous due to the presence of the thoria particles. This made crevice formation (incipient cracking) within slip bands more difficult in TD-nickel than in Ni-200. Both materials cyclically hardened to a constant (saturation) flow stress which increased with increasing plastic strain amplitude. Cellular substructures were developed in both materials during cycling. The cell size in TD-nickel was controlled by the thoria particle distribution and was independent of plastic strain amplitude over the range investigated. The cell size in Ni-ZOO was larger than that in TD-nickel at similar plastic strain amplitudes and was a function of plastic strain amplitude. These results, together with the cyclic stress-strain curves for both materials, are discussed in terms of a model for fatigue strain accommodation at saturation recently proposed by Feltner and Laird. NUMEROUS fatigue investigations have considered the interrelation of slip character, dislocation substructure, and cracking in pure metals and solid-solution alloys. However, except for the studies of the low-strain fatigue of internally oxidized copper alloys1 and cast, dispersion-strengthened lead,&apos; little is known about the effects which small, incoherent, nondeform-able, second-phase particles have on cyclic deformation and cracking processes. Effects due to the particles alone are often obscured by a dislocation substructure introduced during thermomechanical processing of dispersion-strengthened metals. In the present study, recrystallized TD-nickel and a commercially pure nickel, Ni-200, were employed to evaluate the effect of a thoria dispersion on high-strai fatigue deformation and cracking at room temperature. I) MATERIAL AND EXPERIMENTAL PROCEDURE The TD-nickel was supplied by DuPont as a 5/8-in.-thick stress-relieved plate which had been subjected to a proprietary schedule of thermomechanical treatments, and the Ni-200 as 3/4-in. bar which was subsequently annealed for 2 hr at 850°C in argon resulting in an average grain diameter of 0.05 mm. The compositions of these materials are given in Table I. The microstructure of the TD-nickel consisted of elongated grains parallel to the primary working direction with an average width of 0.16 mm, Fig. l(a). Many fine annealing twins were present indicating that the starting material was in a recrystallized condition; this supposition was confirmed by the absence of of any extensive dislocation substructure, Fig. l(b). Sheetlike stringers parallel to the rolling direction were occasionally seen both within grains and at grain boundaries. Some approximately spherical particles about 2 in diam, which may correspond to exceptionally large thoria particle aggregates, were also present. The average Young&apos;s modulus of the plate material in the rolling direction was 21.8 X 106 psi which is consistent with a {100}<001>recrystalliza-tion texture3&apos;* being prominent. In transmission microscopy, the 2.3 vol pct of thoria particles generally appeared to be uniformly distributed although some clusters, 0.1 to 0.3 in diam, of larger particles were observed as previously reported for TD-nickel sheet,5 and stringering of particles was present in some areas as welt. The average diameter of the thoria particles was 450A with a calculated mean planar center-to-center spacing of 2100A, as determined by quantitative metallographic analysis.= The 0.2 pct offset yield stress was 36,000 psi which agrees with the value predicted by the modified Orowan relation7 for edge dislocations bowing between thoria particles of the size and spacing observed in the present investigation. Fig. 2 illustrates the specimen design employed for the axial high-strain fatigue testing. Adapters were screwed onto the threaded portions of each specimen so that testing could be performed in the same manner as that reported for buttonhead specimens.8 Stressing was coincident with the working direction for both materials. The gage section of each specimen was electropolished and lightly etched prior to testing. The total strain was controlled, being varied between zero and a maximum tensile strain ranging from 0.009 to 0.036. In addition to these tests, a circum-ferentially notched TD-nickel specimen was cycled over a total strain range of 0.0075. The same strain

Jan 1, 1969

By G. M. Bernard

The Andacollo gold mine is located near the small town of Andacollo, Chile (population 14,000) in the fourth region. It is approximately 35 miles by paved road from the coastal city of La Serena, and Santiago lies about 300 miles to the south (see map.) Site elevation varies between 3300 feet and 4200 feet. The climate at the site is moderate but arid, with an average annual rainfall of only five inches. Because there is little rain, and the temperature varies only a few degrees (57"F to 75°F); there are no freezing, solution dilution, or pad saturation problems. The climatic conditions are ideal for heap leaching.

Jan 1, 1996

By T. Xue, A. Singhal, D. Wrana, I. Mihaylov

"At Vale's Thompson Nickel Refinery, the copper and arsenic in the nickel electrolyte are removed by precipitation using H2S, generating a sulphide precipitate containing about 40% Cu, 8% Ni, 2 to 4% As and 30% S. This residue has been typically stored in impoundment ponds within the plant site. Various options for reprocessing it have been investigated over the years. One of the options involved oxidative pressure acid leach, whereby over 95% Cu, Ni and As were extracted. More recently, an atmospheric alkaline sulphide leaching process for selective removal of 70 to 90% As was developed and mini-plant tested. After alkaline sulphide leaching, the dissolved arsenic is precipitated as stable basic ferric arsenate (BFA) with the addition of ferric sulfate, while the low-arsenic Cu/Ni sulfide is suitable for processing in Vale's operations elsewhere in Canada.INTRODUCTIONVale’s Thompson Nickel Refinery was commissioned in 1961 and currently it has the capacity to produce about 56,700 tonnes per year of electrolytic nickel as cathode slab and rounds products. A sulphate and chloride mixed electrolyte is used in the Thompson refining process and it typically contains 75 to 85 g/L Ni, 50 to 60 g/L Cl, 90 to 120 g/L SO4 2-, 25 to 30 g/L Na and 8 to 9 g/L boric acid. The main impurities in the Thompson anolyte are copper, cobalt, arsenic, iron and lead.The simplified Thompson electrolyte purification flowsheet is shown in Figure 1 (Oliver & Wrana, 2009, Goble & Chapman, 1986); it consists of precipitation of copper, arsenic and lead with H2S, followed by nickel replenishment with ground matte and acid and removal of cobalt and iron with Cl2 and NiCO3."

Jan 1, 2012

By J. Cayouette, C. Bolduc, P. Blatter

"The LaRonde mill was commissioned in 1988 and began treating ore from the Penna shaft in 2000. The mill has been expanded to accommodate 7,500 t/day, but ramped down to 6,000 t/day since the deep extension began. The process recovers gold, silver, and base metals. In 2008, the Goldex plant began trucking gold-pyrite concentrate to a treatment plant feeding pulp to the LaRonde mill, increasing gold output by 40,000 oz. In 2009, the high-grade Lapa ore processing facilities were annexed to the LaRonde plant, increasing gold output by 120,000 oz. This paper outlines the metallurgical considerations related to the treatment of different ore types.RÉSUMÉ Le concentrateur de la mine LaRonde a été mis en service en 1988 et a commencé à traiter le minerai du puits Penna en 2000. La capacité du concentrateur a augmenté afin de permettre le traitement de 7 500 tonnes de minerai par jour, capacité qui est retombée à 6 000 tonnes par jour (tpj) depuis le début de l’extension en profondeur de la mine. Ce procédé permet de récupérer de l’or, de l’argent et des métaux communs. En 2008, l’usine de traitement Goldex a commencé à transporter par camion du concentré de pyrite et d’or vers une usine de traitement alimentant en pulpe le concentrateur de LaRonde, augmentant ainsi la production d’or de 40 000 onces. En 2009, les installations de traitement du minerai à haute teneur de Lapa ont été annexées à l’usine de traitement de LaRonde, augmentant la production d’or de 120 000 onces. Cet article passe en revue les considérations métallurgiques liées au traitement de différents types de minerai.INTRODUCTIONThe LaRonde metallurgical plant in the Abitibi region of Quebec, Canada, evolved into a mature milling site following expansions in 2000–2003, thanks to improved processes, manpower, and water treatment capabilities that better accommodate the complex constituents of the LaRonde ore. The ore requires a sequence of grinding, copper- lead and zinc flotation, and precious metals leaching, ending with a carbon-in-pulp (CIP) gold recovery. A portion (35%) of the tailings is sent back underground as paste backfill after being treated with the SO2-air cyanide destruction process. The tailings area has a dedicated cyanide destruction and metal precipitation area through which the used water passes before returning as recycled water. A biological water treatment plant was built in 2004 to address water toxicity in the final effluent. The water toxicity was the result of the high sulfide content of the LaRonde ore (from thiocyanate or stable cyanide species buildup) and the use of 90% recirculated water in the mill. The water treatment plant uses bacteria to transform thiocyanate into nitrates and precipitate phosphorus from the water before it is discharged into the environment."

Jan 1, 2017

By Theodore M. Gurr, James J. Bartels

Phosphate produced fertilizers provide the phosphorous nutrients required by plant life to sustain significant growth, thus improving the production of food for the world&apos;s population. Phosphorous replenishment of soil currently can be obtained efficiently by no other means than by direct application of phosphate fertilizer to the soil. Phosphate products are also utilized in animal feeds, detergent, and various industrial processes. Thirty-three countries are presently mining phosphate rock for the production of fertilizers, and its process byproducts. The major phosphate rock producers are the United States, the former Soviet Union, Morocco, and China [(Table 1)]. Phosphate rock is generally defined as a rock material that contains phosphate minerals which are sufficient for commercial usage. Phosphate minerals are found in sedimentary igneous and metamorphic rocks. Economic resources of phosphates are primarily developed from sedimentary rock sources. The apatite mineral family comprises the majority of phosphate constituents from both sedimentary and igneous rocks. Igneous phosphate rocks are generally composed of fluroapatite whereas sedimentary rocks are generally composed of carbonate fluroapatite, or chlorapatite. Phosphate rock is mined by a multitude of processes, including surface and underground mining. The mined rock requires mechanical and chemical processing to liberate the phosphate for utilization as fertilizer. Approximately 85% of the world&apos;s phosphate production uses sulfuric acid, with the remaining processes utilizing nitric and phosphoric acids. Ammonia is also introduced in the manufacture of liquid fertilizer, ammonium phosphate and ammonium polyphosphate. GEOLOGY Mineralogy and Chemical Properties Phosphate most commonly is derived from the mineral apatite, which is chemically described as CA5(F,Cl,OH)(PO4)3. Apatite is also broken up into the composite apatite minerals which are fluroapatite Ca5F(PO4)3, chlorapatite Ca5Cl (PO4)3 and hydroxyl-apatite Ca5(OH)(PO4)3. Carbonate CO3 can substitute for PO, forming carbonate apatite, called francolite. The guano phosphate (bird excrement) mineralogy commonly occurs as brushite CaHP04.2H20, monetite CaHPO4, whitlockite Ca3(PO4), and dahlite Ca10(PO4)6(OH)2. Apatite minerals form from igneous sources where they are developed deep in the earth&apos;s crust; and the cooling processes are slow, producing long, hexagonal prismatic to tabular crystals. The termination of the crystal can be basal plans or pyramids. In some instances, the crystals are bipyramids. Faster cooling, or sedimentary derived apatites are more commonly cryptocrystalline. Collophane is a name given to cryptocrystalline apatite found in phosphate rock in fossil form. Apatite minerals are relatively soft and are used to describe the 5 hardness on the Mohs scale. Cleavage is poor and develops along the C(0001). The color of apatite minerals varies from colorless, to violet, and blue, but are predominantly green or brown; luster is vitreous to subresinous. Collophane&apos;s physical appearance is often opaline, dense, with colloform structure and sometimes concretionary, nodular or pulverulent. Phosphate Bearing Materials Origins Phosphate deposits of igneous and metamorphic origins have been well defined. Sedimentary origins have been defined, but with a greater degree of uncertainty. Research in the 1980s proved out some of these processes through studies of real time deposition in Chile, Peru, and Australia. Igneous intrusive alkali rock and associated contact metamorphic rocks, provide approximately 20% of the world&apos;s phosphate. Fluorine containing apatite minerals are the most common materials containing phosphorous. Fluorapatite occurs in most igneous rocks. Fluorapatite and fluorine hydroxylapatite, together with carbonate varieties of these, are important members of the group. Where an essentially pure chlorapatite, carbonate apatite, and hydroxylapatite are rare and restricted in occurrence, the fluorine containing types occur in most all igneous rocks as early formed accessory mineral, usually in microscopic crystals, and may occur as extremely large bodies as magmatic segregations from alkalic igneous rocks. Apatite is also found crystallized in pegmatitic faces of both acidic and basic types of igneous rocks. More specifically, apatite is associated with magnetite deposits, in hydrothermal veins, especially those formed at relatively high temperatures, and in veins of the Alpine type. Apatite is common in both regionally and contact metamorphosed rocks, especially in the crystalline limestones where it is associated with sphene, zircon, pyroxine amphibole, spinel vesuvianite, phlogopite, talc, chloride schists, and as a contact metamorphic mineral.

Jan 1, 1994

By Frederick McNeill, John Harper, Mahmood Khwaja, Shawn Lavoie, Timothy Cloughert, David Polcari

City of Manchester, NH, is implementing a 20-year plan to control and significantly reduce overflow of its sewer system. The long-term control plan had called for large scale box culvert installation through downtown Manchester. The City and the project team evaluated, and eventually selected, tunneling as an alternative to mitigate construction related impacts for open trench approach. This paper provides an overview of the alternative evaluation process; discusses the development of the preliminary design for the 12,000 feet long, 12-ft internal diameter tunnel, including plan and profile selection, subsurface conditions, drop shafts and the anticipated packaging of the separation areas.

Dec 1, 2022

By José Fernandes De Oliveira Ugarte

El presente texto analiza las diseminaciones de oro a través de la microscopia electrónica de barrido con microanálisis dispersivo. Dicha metodología permite identificar el oro refractario diseminado ( forma nativa y "electrum") en diferentes ambientes mineralógicos. Asimismo, genera informaciones acerca de la composición química, tamaño de partículas y grado de liberación del mineral de valor económico.

Jan 1, 2013

By Edinson Terán, Delia Cieza

El Perú, enfrenta crisis y entrampamiento político institucional; sin embargo, las regiones evidencian diversos esfuerzos por superar las limitaciones políticas, mediante la participación del conjunto de actores en la gestión de su desarrollo. La presente experiencia parte de este proceso, Cajamarca tiene un especial significado considerando que se convirtió en el icono de las manifestaciones en contra de la minería en el 2011. Actualmente la región vive procesos de cambio con iniciativas de diversos actores que buscan nuevas condiciones para aprovechar sus recursos y convertirla en un territorio competitivo nacional e internacionalmente.

Sep 29, 2023