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By Naum S. Malts, Victor M. Sizyakov, Nikolai S. Shmorgunenko

A method has been developed in the USSR for processing the low-grade high-silica raw material by sintering the aluminacontaining mixes, which is widely used both as a principal process and in combination with the Bayer process. The process parameters were obtained on the basis of a study of interaction and equilibrium states of the individual profiles of a multi-component system Na20-A1203-Fe203-CaO-Si02 at the sintering temperatures. The compositions and melting temperatures of eutectics of the individual profiles are given. It is shown that the most easily meltable mix is one having a red mud as the main component. The variations are given of the sinter physical characteristics and recovery of the useful components with the sintering temperature. Considerations are given on selection of a sinter leaching flowsheet and directions are shown for improving a series method of processing the low-grade raw material.

Jan 1, 1984

By Claude Bazin

The Quebec province in Canada is a major producer of aluminium through electrolysis of alumina. The alumina that feeds the electrolysis cells is mainly imported from foreign countries, where it is extracted from bauxite. Other alumina sources, such as clays, have sometimes been considered as potential substitutes for bauxite. There is a large clay deposit in the area of Murdochville in Quebec, with sulfuric acid and electricity available nearby the orebody, as well as sea port facilities leading to the major aluminium plants in Quebec. These are favorable factors to assess the economics of an industrial plant to produce alumina from clays. This paper reviews some of the processes considered for the extraction of alumina from clays and presents results of the initial work carried out on the clay available from Murdochville.

May 1, 2001

By R. S. Dean

JUST as the mineralogical name bauxite has come to include several minerals not known at the time the name was first applied, so the concept of bauxite as the one source of alumina must be enlarged during the war period to include the production of alumina from other, low-grade sources. Speaking in a narrow mineral logical sense, there is no such thing as bauxite, and similarly there is now no one source of alumina, and no one process for its extraction.

Jan 1, 1943

By G. M. Mason

History Due to its strong affinity for oxygen, aluminium as an element was isolated only after many other rarer metals had been prepared and studied. It was just over one hundred years ago that two scientists, Hans Christian Oersted and Friedrich Wohler, produced the first few pellets of aluminium and determined its physical properties. These early investigators used metallic potassium as a reducing agent. Later, St. Claire Deville, by using sodium instead of potassium, improved the yield in the reduction and produced some pounds of aluminium. From a figure of about $110 per pound in 1855, the price had fallen to about $4 per pound by 1886. In that year, a very important discovery was made independently by Charles Martin Hall in Oberlin , Ohio, and Paul . Heroult in France. Stated in its simplest terms, the discovery of Hall and Heroult was that aluminium could be produced by electrolyzing a solution of the oxide in molten cryolite. The application of the Hall-Heroult process made the metal immediately available in commercial quantities and, by 1900, the price had fallen to 23 cents per pound. The growth of the industry in the past fifty years is well known, and is outside the scope of this paper. It is, perhaps, not generally known that aluminium was first produced in Canada at Shawinigan Falls in 1901. Since that time, production has been continuous at one or more smelters in the Province of Quebec. The largest of these smelters, which is also the largest aluminium smelter in the world, is at Arvida, Quebec, on the Saguenay river. A large power potential and access to ocean shipping were the two principal reasons for the establishment of aluminium production facilities in the Saguenay district. It is interesting to note that the aluminium industry, as a large user of power, took an active part in the early industrial development at Niagara Falls and Shawinigan Falls, and is now repeating the process in the Saguenay valley.

Jan 1, 1949

By A. W. JR. Whitaker

Introduction Within the last quarter of a century the Lake St. John-Saguenay valley has been converted from a quiet agricultural district to one of the important industrial centres in the Dominion. First there was the expansion -of the pulp and paper industry, and then, starting a little over ten years, ago, the development, in large .blocks, of the enormous hydro-electric power resources of the Saguenay river. Coincident with this power development, the world aluminum industry was rapidly expanding and looking for a new source of cheap power. This was found in the Saguenay valley-and so, in 1925, Arvida was established by the Aluminum Company of Canada as a location for one of the largest aluminum plants in the world. Arvida, an incorporated city comprising an area of some ten square miles, is ideally located on a more or less level plateau about 350 feet above the Saguenay river. Seldom has industry stepped boldly forth and carved out of Nature a community with full city status. The process usually goes through the steps of camp, village, town, and city, involving improper planning or lack of planning; all of which has been avoided at Arvida by starting out in the right way. While this community has only a population of approximately 2,000 at the present time, it is carefully planned for a population many times this figure. Streets, city parks, business and residential sections, have been definitely planned, and, with a City government, proper ordinances were passed at the beginning to ensure the building and maintenance of a model industrial community. As the crow flies, Arvida is about 35 miles due east of lake St. John, and 0 miles west of the St. Lawrence river at Tadoussac. From a shipping standpoint, it is well served by the Canadian National railways, with direct lines to Quebec and Montreal, and by ocean shipping facilities at Port Alfred, which is about 15 miles to the east on the Saguenay river and is served by the Roberval & Saguenay railway. The latter has interchange connections with the Canadian National at Arivda, and also operates a branch line into the city of Chicoutimi.

Jan 1, 1933

By Norman Ketzlach

Manganese Products, Inc. has developed a chemical process for the recovery of high-grade manganese oxides from low-grade manganese ores. Ammonium sulphate is also produced. Manganese ore is leached with sulphurous-sulphuric acids. The impurities are removed with ammonia, heat, and air. Manganese oxides are produced by addition of more ammonia and air under pressure. Ammonium sulphate is recovered from the remaining solution.

Jan 3, 1950

By F. Garcia-Carcedo

This research aimed to develop a clean technology for the production of ZnO from residues generated by the steel making industry. Scientific and technologic parameters were both investigated in order to define an industrial methodology allowing the integral recycling of EAF dusts. A reengineering of the Waelz process was perform,~d, aiming to produce a clean fraction of ZnO, fluorine free, as well as a slag enriched in metallic iron, which can be recycled in electric arc furnaces. The technical conditions used to produce the ultra pure fraction of ZnO are described. The magnetic separation of the EAF dust allowing the production of a phase high in iron and recyclable to the electric arc furnace is also discussed. Thermodynamic of the volatilization of all the elements contained in the EAF dust was performed in order to better control the thermal process in the rotary kiln. Tests were run using different reducing agents, weakly or highly reactive, aiming to determine their specific efficiency.

Jan 1, 2000

By I Wilkomirsky

The neutralization of S02 from copper smelter gases to produce calcium sulphate (anhydrite) is considered an attractive alternative to sulphuric acid production. To develop the required technology for the process, a three-stages fluidized bed reactor prototype was designed and operated using both high and low grade limestone and gases with 2, 8 and 12 vol -% S02. The results obtained shows that at 800°C and up to 8 vol-% S02 in the gases, the capture of S02 reaches near 99% with limestone conversion to anhydrite of 45 to 48%. A larger pilot plant is presently being built at the Codelco El Teniente-Caletones Smelter to demostrate the technology and to evaluate the economics of the process. Cu2007

Jan 1, 2007

By John A. Oberteuffer, James E. Lawver, Ionel Wechsler, Bo Arvidson

INTRODUCTION The demand for aluminum metal and thus for alumina and bauxite (the primary sources), is expected to increase at an annual rate of about 6% through the year 1980. It appears that about 90% of the future aluminum metal requirements could be produced at domestic plants but the production would remain dependent on the use of imported bauxite and alumina. The U.S. dependence on foreign bauxite and alumina is of considerable economic and strategic concern because availability and prices are very sensitive to changes in the political climate in the exporting countries. Dependence of the United States on foreign sources of bauxite closely parallels its dependence on foreign sources of oil, particularly since ten countries, including most of the world's major bauxite exporting countries, recently have formed the Inter- national Bauxite Association (IBA) for the specific purpose of increasing revenue for the participating countries. Although it is clear that long-term aluminum requirements cannot be fulfilled by domestic resources of bauxite,; the U.S. has an almost unlimited supply of aluminous resources such as Kaolinite (38% Al203), Alunite (37% Al203), Dawsonite (35% A1203), and Anorthosite (29% A1203)

Jan 1, 1976

By Tugba Basargan Çagatay Morall, Gülhayat Nasun-Saygili

"The aim of this study is to prepare the apatitic material using colemanite mineral (2CaO.3B2O3.5H2O) which was obtained from Eastern Anatolia, Turkey. For this purpose, colemanite mineral was reacted with different phosphate sources, namely, dipotassiumhydrogen phosphate, ammoniumhydrogen phosphate and ortophosphoric acid at various temperatures and time periods. Experiments were run in batch system. The apatitic material was prepared by a wet method and followed by heat treatment at various temperatures. Structure of the samples was analyzed by XRD, FTIR, XRF and SEM. The particle size analysis was also made for the result product. The experiments showed that reaction temperature, contact time and heating temperature are important parameters to transform the colemanite into apatitic structure. Reaction of colemanite with dipotassiumhydrogen phosphate and ammoniumhydrogen phosphate results in amorphous, non-stoichiometric hydroxyapatite type structure, while acid based phosphate source gives brushite and monetite type product before and after heat treatment, respectively.IntroductionBoron which has properties intermediate between metals and non-metals is a member of semiconductor group of elements. Its atom is very small and has three valence electrons. Boron exists in Earth's crust (from 5 mg/kg in basalts to 100 mg/kg in shale) and in the oceans (4.5 mg/l) [1]. Turkey has 60% of world's boron reserves which is mostly in the form of colemanite (2Ca0.3B203.5H20). Colemanite is the most important calcium containing commercial borate mineral with 5 mol of crystal water [2].Calcium phosphate cements are very useful for bone repair applications because of its chemical structure. They are injectable, easy to shape and can be maintained locally. Because of those reasons calcium phosphate cements are very effective to fill bone defects with an irregular shape [3]. Hydroxyapatite (Ca10(PO)6(OHh) is the major component of the bone (70 wt%). However, hydroxyapatite stoichiometry, composition and crystallinity differ in human skeleton system [4, 5].The main objective of the present work was to study the transformation of colemanite to apatitic material using different phosphate solutions at several time spans and temperatures. The experiments were run in a batch system, change in pH and also weight loss were measured after the experiments. Several analysis methods were used to investigate the chemical and structural changes."

Jan 1, 2012

By F. M. Isik

Due to not to being our own sources, some energy sources such as oil and natural gas which are used frequently, are not reliable energy sources for our country. Coal is an important energy source that Turkey also has a certain reserve as many countries have. To be able to use for many years the reserves of this source that comes from fossils, there is a great need to make more investigation and development activities on it. On behalf of using more productively the current sources, briquetting facility is represented that commenced production in The Aegean Lignite Establishment in 2005 which is in Manisa, Soma. This facility is a part of General Directorate of Turkish Coal. This facility is run by Soma Coal Establishment that works in return for royalty in a licensed field that belongs to TKI. As the demand of market for 0.5 mm dust coal diminishes, we had to accumulate and stock unmerchantable dust coal. Besides, we have intermixed 0.5- 1 mm powder coal and 0.1 ? 0.5 mm fine powder coal obtained as a result of washing in coal washer with thermic coal with low price as we could not find a client in the market. In this study, the production stages of briquette coal with high quality produced as a result of pressing by mixing 0.5- 1 mm powder coal and 0.1 ? 0.5 mm fine powder coal with molasse and lime, quality control, and data?s obtained about calories and solidity of products were presented. By briquetting of 0.5- 1 mm powder coal and 0.1 ? 0.5 mm fine powder coal washed in coal washer, it became possible to sell these coals on a high unit price and by providing the demand of coal of the market, we contributed to productivity of our firm. Keywords: briquette coal, briquetting, producing stage

Sep 1, 2012

By Bowmaker MD

Zinc and lead sulphide flotation con centrates containing from 0.02 to 0.20% Cd are roasted on an updraught sinter machine at the Cockle Creek Works of Sulphide Corp- oration Pty. Limited to provide sinter for simultaneous smelting of zinc and lead in an Imperial Smelting Furnace (I.S.F.). The major portion (about 85%) of cad- mium in concentrate feed to the Sinter Plant is eliminated with sinter fume and is rec- overed by weak acid leaching in the Cadmium Plant'followed by counter-current decantat- ion, ion exchange concentration, precipitat- ion of cadmium sponge and partial refining to caustic refined cadmium metal. Cadmium not eliminated during sintering is volatilised in the I.S.F. and subsequently recovered as zinc alloy by distillation in the zinc ref- ining columns of the Zinc Refluxer. The partially refined cadmium metal and cadmium- zinc alloy are blended as feed to the cadm- ium refining column where separation and recovery of cadmium is effected. Condensate from the column is further refined under caustic soda flux and cast into anode shapes. Production is 200 t.p.a. at 85% recovery of cadmium in concentrates. Emphasis is placed on changes from the practice described by Baker and Munro (1965) and upon operation

Jan 1, 1972

By Fred W. Menzl, Raymond E. Sutton

GABBS, Nev., in Nye County, is 140 miles southeast of Reno and about 31 miles north of the Southern Pacific railroad at Luning. The townsite lies at the eastern edge of Gabbs Valley, and plant and quarry operations are in the foothills of the Paradise Range at elevations from 5000 to 5900 ft. Hard-surfaced roads connect with main U. S. high- ways on the north and south. Historical Background: Brucite, a magnesium hydroxide mineral, was discovered in the Gabbs area in 1927 and magnesite, the magnesium carbonate, was found later on claims originally staked and worked for the tungsten mineral scheelite. The brucite is on a granodiorite contact and the magnesite is usually some distance away from the granodiorite

Jan 7, 1957

By Ahmad-Bitar¸ Riyad N., Shadia J. Ikhmayies

"CdSi-xTex thin films with O<x<0.3 were prepared on glass substrates by the annealing of CdS:In thin films produced by the spray pyrolysis technique. The annealing was performed in nitrogen atmosphere at 400 --,C in the presence of Te vapor. The compositions of the films were determined by energy dispersive X-ray detection (EDAX) measurements. Scanning electron microscopy (SEM) was used to investigate the morphology of the films. The transmittance of the films was measured at room temperature in the wavelength range 400-1100 nm and used to produce the absorbance of the films. The first derivative of the absorbance was used to estimate the optical bandgap energy Eg by using the minima in the first derivative. The relation between the bandgap energy and the fraction x was found to fit with a parabola and bandgaps in the range 2.196: =Eg: =2.480 e V were obtained. Urbach tailing in the bandgap was also investigated.IntroductionPolycrystalline thin film CdS/CdTe solar cells are promising candidates for low cost and high efficiency photovoltaic applications. The fabrication process of these cells involves the deposition of a p-type CdTe absorber layer on top of an n-type CdS one which acts as a window. The efficiency of the solar cell is affected by the method of preparation of both layers [1, 2].Interdiffusion processes taking place at the CdS/CdTe interface have largely been realized to significantly affect the device performance [3] by changing the bandgap of both CdTe and CdS which modifies the spectral response of the solar cell [ 4]. The formation of a mixed crystal or a graded layer causes a reduction in the lattice mismatch at the CdS/CdTe junction [5] which is about 9% [ 6] and then a reduction in the number of interfacial states and recombination centers, thus enhancing the solar cell performance [7, 8]. A complete understanding of the properties of the intermixed layers CdSxTe1-x that usually constitutes the interfacial region in the CdS/CdTe solar cell is essential. According to our knowledge, a lot of research was devoted to the investigation of the diffusion of S into CdTe [3-5, 9]; but little on the direct investigation of the diffusion of Te into CdS [6]."

Jan 1, 2011

By J. Y. Lee, J. Kang

"A selective chlorination process using cuprous chloride (CuCl) was investigated for producing high-grade chalcocite (Cu2S) from chalcopyrite (CuFeS2) feed. The chalcopyrite feed was reacted with CuCl in the presence or absence of sulfur (S) at temperatures between 900 and 1,100 K for five hours under an argon gas atmosphere. The experimental results indicate that iron (Fe) was selectively removed from the chalcopyrite feed as ferrous chloride (FeCl2) at 1,100K and the mass percentage of Fe thus decreased from 31.5 percent in the feed to less than 0.05 percent in the residue, under certain conditions. As a result, high-grade Cu2S was obtained from chalcopyrite in a single step. This new selective chlorination of chalcopyrite using CuCl as a chlorinating agent is therefore feasible for producing high-grade Cu2S.IntroductionCurrently, blister copper (Cu) is commercially produced from sulfide ores by pyrometallurgical methods. These methods entail matte smelting and converting, with the removal of iron (Fe) and sulfur (S) as slag — mainly Fe2SiO4 — and sulfur dioxide (SO2) gas, respectively.In the pyrometallurgical method, as shown in Fig. 1 (Schlesinger et al., 2011), Fe is removed through two steps at high temperatures of around 1,500 K to reduce the loss of Cu during slag formation. However, the slag still contains 1 to 8 percent Cu. Therefore, Cu in the slag is recovered by slag treatment processes, after which the concentration of Cu in the slag decreases to around 0.4 to 1.3 percent. Although Cu is recovered, the slag that had undergone the treatment processes is difficult to use as a feedstock for ironmaking owing to the small amount of Cu remaining in the slag (Schlesinger et al., 2011).In addition, large amounts of SO2 gas are generated by the oxidation of S in the current Cu production process. The SO2 gas generated is captured and used to produce sulfuric acid (H2SO4). However, despite the development of SO2 gas-capture technologies, the emission of SO2 gas is still a problem in terms of environmental considerations.Several studies had been carried out in the past to overcome the drawbacks of the current process. The extraction of Cu from sulfide ores using a chlorination process is one alternative that had been investigated by several researchers. Most chlorination processes used chlorine gas (Cl2) as a chlorinating agent at temperatures between 573 and 1,023 K under an inert atmosphere (Donaldson and Kershner, 1962; Olsen et al., 1973; Kumar, Li and Warren, 1985, 1987; Kanari et al., 1999; Kanari, Gaballah and Allain, 2001)."

Jan 1, 2017

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Jan 1, 1942

In, one stage of the purification sequence of the zinc sulphate electrolyte used at Risdon, cobalt is, precipitated as cobalt nitroso-,B-naphtholate, which is separated from the solution by filtration. Owing to the method of conducting the purification the filter cake contains much gypsum, unconsumed calcium hydroxide, basic zinc sulphate, and entrained zinc sulphate solution, Small amounts of iron and copper that have escaped previous purification steps are also precipitated with the cobalt as nitroso-,B-naphtholate. Thisprecipitate, normally carrying about 18% cobalt, is the raw material for the production of cobalt oxide, cobalt sulphate and cobalt metal.COBALT OXIDEThe main problem in the preparation of a reasonable grade of cobalt oxide from the filter cake is the economic separation ,of the cobalt nitroso-,B-naphtholate from the gypsum. This was solved by the application of flotation. It was discovered that the cobalt nitroso-,B-naphtholate, under correct conditions, would float without any addition agent, giving a relatively high grade concentrate. A plant, was built and has been in operation since 1936. The filter cake is pulped with water and pumped to one of two mechanically agitated tanks, eath capable of holding one day&apos;s output. Sulphuric acid is added until the free acid concentration reaches about 3 gm. per litre. Basic zinc sulphate is dissolved and calcium hydroxide converted to gypsum. The pulp, carrying about 2t lb. of solids per gallon, is pumped at a suitable rate to the first of a series of 10 flotation, boxes.. Each box is 3 ft. by 2 ft. 9 in. by 5 ft. deep, and, all have a common, pulp level, since the...

Jan 1, 1949

Early data about the amounts of coal used in making coke, and even of the first oven plants are meager. Table 84 shows the data about beehive ovens from 1841 to 1940, inclusive, and Table 85 shows the data about byproduct ovens from 1895 to 1940, inclusive.

Jan 1, 1942

By B. S. Ignasiak, Dennis Carson

"Laboratory study of the behaviour• of binary coal blends shows that coke strength is more closely dependent on the distribution of ""reactive"" oxygen in the parent coals than on their proximate, elemental or petrographic compositions, and it also appears possible to qualitatively predict coke strength from information about reactive oxygen. Proceeding from this recognition, experiments were carried out to test the possibility of preparing metallurgical cokes from binary blends in which non-caking hvb coal was substantially freed of reactive oxygen by prior heat t1•eatment under controlled conditions. The results indicate that such blends can yield surprisingly strong cokes and that coke strength may be amenable to further improvements by modifying the size distribution of the pre-heated component and the rate at which it is charred."

Jan 1, 1976

By P. R. Smith, R. A. Walters

Borates have been mined in the desert areas of California and Nevada for more than 100 years. To about 1890, playa surface mining provided the chief sources of boron minerals. Underground mining of colemanite and later of borax and kernite was predominate until about twenty years ago. Open pit mining of the large deposits of borax and kernite near Boron, California has been most significant for the past twenty years. Mining of colemanite in the Ryan, California area, near Death Valley, began in 1907. Following the discovery of the large deposits in the Boron area (about 1957), mining in the Death Valley area became nearly nonexistent. Only small tonnages were mined for special uses. Little mining was done in the Boraxo area near Ryan. The first claim was made in about 1915. In 1960 the area became the property of the Kern County Land Company, which was acquired by Tenneco Oil Company in 1967. In 1976 the various boiate properties and claims in this region were acquired by American Borate Corporation. The open pit mine is now approximately 122 m (400 ft) deep, 910 m (3000 ft) long and 305 m (1000 ft wide). The borates in the Boraxo pit consist primarily of three minerals. These are about 50% colemanite (CB6011 5H20), about 40% probertite (NaCaB50g 5H20), and 10% ulexite (NaCaBgOg 5H20). The colemanite, along with boric acid and high-grade colemanite ore from Turkey provide the only sodium-free borates for production of textile grade fiberglas. When heated to its decomposition temperature, colemanite decrepitates to a fine powder, which is the basis for the concentration process. The gangue minerals in this deposit are primarily calcite and clays, including bentonite. The ore body has a very low arsenic content, which is a desirable feature. Test work had been done with samples prior to the results discussed herein. This paper will discuss results of test work which were the basis for erection of a plant, and the subsequent plant operations. Laboratory Calcination and Air Tabling Tests Laboratory calcination tests showed that substantial upgrading of the borate could be accomplished by calcining followed by screening of the calcined material. Removal of the + 28 mesh calcine resulted in borate losses of less than 10% with a rejection of 40 weight % or more of the calcine. The minus 65 mesh calcine generally met the requirement of containing 48%, or more, B203. The minus 28 plus 65 mesh material contained an intermediate quantity of borate and would require additional treatment. Testing demonstrated that ore would not have to be reduced to a size finer than 19 mm (3/4 in.) prior to calcination. A temperature range from 400 to 455OC (750 to 850°F) was apparently satisfactory. Calcination at a temperature of 48Z°C (900°F), or higher, was unsatisfactory due to fusion. All laboratory calcination tests were static tests conducted by placing small covered charges in a laboratory furnace for 40 min. In all tests vapor issues from the furnace for 5 to 7 min. Following this period the ore could be heard "popping," due to decrepitation of the colemanite. The reaction generally continued for approximately one-half hour. Various size fractions of the calcination products from laboratory tests were subjected to laboratory air tabling tests, usually after removing the plus 28 mesh material. Laboratory air tabling tests were conducted employing a Whippet V-80 model air table manufactured by Sutton, Steele and Steele Co. now known as Tripple S Dynamics. Variables include both end and side-tilt, speed of vibration, and quantity of air rising through the deck. In addition to the variables in the machine itself, the feed rate is also a rather critical variable. Testing demonstrated that all - 28 mesh size fractions of the calcine could be successfully concentrated to 48% F2O3 or greater. For the finer material recoveries into the concentrate were between 85 and 90% of the borate. With the coarser material a substantial amount of middling was produced which required cleaner tabling. Laboratory calcination and air tabling tests indicated a process whereby the borate could be concentrated to about 50% B203 with borate recoveries approaching 90%. Moreover, the iron content of the concentrate was well below the required specification of 0.3% Fe2O3. Pilot Plant Calcination Following the laboratory test work described above, pilot plant testing was conducted to prove the process, provide data for engineering studies, and provide product for a prospective purchaser. The kiln used was 0.9 m-diam (3 ft) by 9.0 m (30 ft) long and had a belly section 1.2 m-diam (4 ft) by 2.74 m (9 ft) long near the discharge end. The kiln was operated at a speed of 0.7 rpm. Gas was fired into the kiln at an average rate of 27.1 m3/hr (958.4 cu ft per hr). The air to gas ratio used was 10:1. The ore was fed to the kiln countercurrent to the flame and discharged through a hopper into a screw conveyor which discharged to a 1.2 m (48 in.) Sweco separator. The separator had 28, 65, and 150 mesh screen cloths, with the plus 28 mesh fraction being discarded. The minus 28 mesh fractions were later subjected to air tabling. The exit gases, containing some calcine dust, were swept through two cyclones to recover the dust. The gases then were scrubbed in a Ducon scrubber; very little dust reported past the first cyclone. The dust from the first cyclone was also saved in drums. In addition to the gas rate, the flue gas velocity, after

Jan 1, 1981