A simple and efficient method for recycling tungsten carbide waste

Abstract

This invention provides a simple and efficient method for recycling tungsten carbide waste containing cobalt and other doped metals (more than 8%). The specific steps include: oxidizing the tungsten carbide (containing cobalt) waste at high temperature into oxides and tungstates; subjecting it to high-temperature alkaline leaching to form cobalt and other metal elements as a slag phase; and dissolving tungsten in sodium tungstate solution, thus achieving solid-liquid separation. The sodium tungstate solution is then converted into ammonium tungstate solution through an extraction process, followed by evaporation and crystallization to form ammonium paratungstate. The method provided by this invention is simple to operate, uses inexpensive and readily available raw materials, and has low processing difficulty. Furthermore, tungsten, cobalt, and other valuable metals can be almost completely separated and recovered, with recovery rates exceeding 98%, resulting in extremely high economic benefits.

Description

Technical Field

[0001] This invention relates to the field of tungsten waste recycling technology, and in particular to a simple and efficient method for recycling cemented carbide tungsten carbide waste. Background Technology

[0002] Tungsten possesses extremely high hardness and strength, and is widely used in the synthesis of cemented carbide, making it an important strategic resource. With the increasing exploitation of tungsten ore, tungsten resources are becoming scarce, and the price of tungsten cemented carbide is high. Therefore, the secondary recycling of tungsten cemented carbide resources is particularly important. However, since tungsten waste is mainly composed of cemented carbide, its processing is difficult. Existing technologies for tungsten waste recycling include electrolysis, zinc smelting, nitrate smelting, and roasting and alkali leaching.

[0003] Current technologies have various drawbacks. For example, the zinc smelting method is only suitable for processing cemented carbides with a cobalt content of less than 10%, has high power consumption, requires sophisticated zinc vapor recovery equipment, and causes significant zinc volatilization pollution. The nitrate smelting method has a long industrial process, uses expensive raw and auxiliary materials, has high production costs, and causes environmental pollution from exhaust gases; moreover, it generates a lot of wastewater and suffers significant losses of raw and auxiliary materials during subsequent product production.

[0004] Currently, conventional roasting and alkali leaching processes result in severe furnace slagging in the roasting equipment, high tungsten content in the slag, low recovery rate, and high cost of secondary slag treatment. Furthermore, the subsequent production of products through traditional metallurgical processes generates a large amount of wastewater and significant losses of raw and auxiliary materials. Summary of the Invention

[0005] To overcome the shortcomings of existing technologies, this invention provides a simple and efficient method for recycling cemented carbide tungsten carbide waste.

[0006] To address the problems existing in the prior art, this invention provides a simple and efficient method for recycling cemented carbide tungsten carbide waste. The method is characterized by two main steps: calcination and leaching of the tungsten carbide waste, and extraction and crystallization with sodium tungstate solution to prepare ammonium paratungstate. This method is applicable to the efficient recycling of all transition metal modified tungsten waste.

[0007] This invention provides a simple and efficient method for recycling tungsten carbide waste, which includes the following steps:

[0008] (1) Prepare tungsten carbide waste and calcine it at different temperatures to destroy its structure;

[0009] (2) Grind the completely calcined tungsten waste into fine powder;

[0010] (3) Add alkaline solution to the calcined and ground fine powder, leach it, and then separate the solid and liquid to obtain sodium tungstate filtrate and filter residue.

[0011] Preferably, the calcination temperature in step (1) is 600-1500℃, the calcination time is 6-48h, and the calcination atmosphere is an oxygen-containing atmosphere. More preferably, the calcination temperature is 750-1000℃, and the calcination time is 10-12h.

[0012] Preferably, the alkaline solution in step (3) is a sodium hydroxide solution with a concentration of 20-46%, a solid-liquid ratio of 1:1-10, a leaching temperature of 60-100℃, and a leaching time of 1-12h. More preferably, the sodium hydroxide solution has a concentration of 25-40%, a solid-liquid ratio of 1:2-8, a leaching temperature of 75-85℃, and a leaching time of 6-8h.

[0013] Preferably, after the leaching is completed, ammonium paratungstate is prepared by extraction and crystallization with sodium tungstate solution, specifically including the following steps:

[0014] (1) Preparation of organic phase: First, acidify N235 with dilute sulfuric acid, then take 2-octanol and kerosene to prepare the organic phase;

[0015] (2) Extraction: Acidify the sodium tungstate filtrate, mix the acidified solution with the organic phase thoroughly and let it stand. After standing and separating the layers, separate the aqueous phase and retain the negative tungsten organic phase.

[0016] (3) Back-extraction: The negative tungsten organic phase is back-extracted with ammonia water, so that the tungsten anions in the organic phase enter the aqueous phase in the form of ammonium tungstate, thereby achieving the purpose of converting sodium tungstate into ammonium tungstate and separating it from the organic phase;

[0017] (4) Evaporation and crystallization: The ammonium tungstate solution after back-extraction is evaporated and crystallized to obtain ammonium paratungstate.

[0018] Preferably, the concentration of the dilute sulfuric acid in step (1) is 5%-35%, and the proportions of N235, 2-octanol, and kerosene are 5-10%, 10-20%, and 70-85%, respectively, with a total of 100%.

[0019] Preferably, the solution in step (2) is acidified to a pH of 1.0-5.0, and the ratio of the aqueous phase to the organic phase in the extraction is 1:1-6. More preferably, the solution is acidified to a pH of 2.5-4.5, and the ratio of the aqueous phase to the organic phase in the extraction is 1:3-6.

[0020] Preferably, the concentration of the ammonia water in step (3) is 10%-28%, the ratio of the ammonia water to the organic matter is 1-6:1, and the evaporation temperature in step (4) is 70-150℃. Preferably, the evaporation temperature is 80-120℃.

[0021] Preferably, the concentration of the back-extraction ammonia is 15%-28%, and the ratio of back-extraction ammonia to organic matter is 3-6:1.

[0022] As another aspect of the present invention, the present invention provides a method for preparing lithium cobalt oxide cathode material from filter residue obtained by tungsten carbide treatment, which includes mixing a lithium source with the filter residue and then sintering it, wherein the molar ratio of the lithium source to the metal in the filter residue is Li:TM=(1.03-1.25):1.

[0023] Preferably, the lithium source is one or more of lithium hydroxide, lithium carbonate, and lithium nitrate.

[0024] Preferably, the sintering process involves a first-stage calcination at 450-600℃ for 4-8 hours, followed by a second-stage calcination at 850-1000℃ for 9-16 hours. The calcination atmosphere in these steps is one of air, oxygen-enriched air, or pure oxygen.

[0025] Furthermore, the filter residue is cobalt hydroxide, which can be directly used in the preparation of lithium cobalt oxide cathode materials or other materials. Other trace elements can be used as doping and modifying elements for lithium cobalt oxide to improve the electrochemical performance of the cathode material. The process includes the following steps: weighing a certain proportion of lithium source, mixing the lithium source with the filter residue evenly, and sintering at high temperature to obtain the lithium cobalt oxide cathode material. Further, in the calcination step of the tungsten carbide waste, the fine powder particle size of the completely calcined tungsten waste after grinding is 1μm-150μm. Preferably, the particle size is 1μm-50μm.

[0026] Further, the lithium source is one or more of lithium hydroxide, lithium carbonate, and lithium nitrate; the ratio of the lithium source to the precursor is Li:TM = (1.03-1.25):1, which is the molar ratio of lithium to metal. The calcination process in the steps consists of a first-stage calcination temperature of 450-600℃ and a holding time of 4-8 hours; and a second-stage calcination temperature of 850-1000℃ and a holding time of 9-16 hours. The calcination atmosphere in the steps is one of air, oxygen-enriched air, or pure oxygen.

[0027] Compared with the prior art, the present invention has the following beneficial effects:

[0028] 1. Unlike traditional alkaline leaching of tungsten ore, which does not require temperatures above 100°C and can completely leach tungsten into the solution under normal pressure, this invention employs a method of calcination followed by alkaline leaching. Calcination reduces the hardness of the hard tungsten carbide (containing cobalt) waste, facilitating powdering, and then complete leaching of tungsten can be achieved through alkaline leaching under normal pressure.

[0029] 2. The entire process is greatly simplified, requiring only two steps: calcination and leaching, to achieve complete separation of tungsten and cobalt, with recovery rates of over 98% for both tungsten and cobalt.

[0030] 3. High-purity ammonium paratungstate can be obtained from the leaching solution of sodium tungstate through two methods: the first is extraction with an organic solvent, and the other is to add acid to precipitate tungstic acid, then dissolve it in hot ammonia water to obtain an ammonium tungstate solution, and then evaporate and crystallize to obtain ammonium paratungstate. Attached Figure Description

[0031] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0032] Figure 1 The flowchart of a method for recycling tungsten carbide waste provided in Embodiment 1 of the present invention is shown.

[0033] Figure 2 The diagram shows the powder of calcined waste material in Embodiment 1 of the present invention, wherein Figure A is a diagram of the hard state of the waste material before calcination, Figure B is a diagram of the state of the waste material after calcination, and Figure C is a diagram of the state of the fine powder obtained by grinding after calcination.

[0034] Figure 3 The XRD pattern of calcination waste powder from Example 1 of this invention is shown.

[0035] Figure 4 The XRD pattern of the leaching residue from Example 1 of the present invention is shown.

[0036] Table 1 shows the W extraction rate under different extractants according to the present invention;

[0037] Table 2 shows the W / Co leaching rates under different leaching conditions according to the present invention. Detailed Implementation

[0038] To facilitate understanding of the present invention, the invention will be described more fully and in detail below with reference to the accompanying drawings and preferred embodiments, but the scope of protection of the present invention is not limited to the following specific embodiments.

[0039] The composition of the tungsten carbide raw material (including cobalt waste) used in this invention is as follows:

[0040] Tungsten carbide raw material composition table

[0041] Element WC Co Ta Cr Ti content / % 91.2 8.7 0.04 0.02 0.04

[0042] Example 1

[0043] Prepare 30g of tungsten carbide (containing cobalt) waste and calcine it in an oxygen tube furnace at 900℃ for 6 hours. Grind it into fine powder using a ball mill (ordinary ball mill, speed 100-300 rpm, time 10-30 min), controlling the particle size of the fine powder to be 1μm-150μm. Add 25% sodium hydroxide solution to the fine powder and leach it at 80℃ for about 4 hours, with a solid-liquid ratio of 1:4. After leaching, separate the cobalt hydroxide slag and sodium tungstate solution.

[0044] The organic phase for extraction was prepared using N235:2-octanol:kerosene in a ratio of 9:11:80. First, 45 ml of N235 was acidified with 30% dilute sulfuric acid. Then, 55 ml of2-octanol and 400 ml of kerosene were used to prepare 500 ml of the organic phase. The filtrate was acidified to pH 3.5 with dilute sulfuric acid. 50 mL of the filtrate and 100 mL of the organic phase were placed in a separatory funnel, shaken for 2-3 minutes, and allowed to stand. After separation, the aqueous phase was collected and retained. The negative tungsten organic phase was back-extracted with 100 ml of 28% ammonia solution. After thorough mixing, the phase was allowed to separate. The aqueous phase was collected, and the ammonia back-extraction process was repeated twice to obtain an ammonium tungstate solution. After removing the oil from the solution, it was evaporated at 75°C to crystallize, yielding ammonium paratungstate crystals. The results of crystal composition analysis and raffinate analysis are shown in the table below.

[0045] Example 1: Composition table of raffinate

[0046] Element W Co Ta Cr Ti Content / ppm 0.9 0.1 / 0.03 /

[0047] Example 1 Crystal Composition Analysis Table

[0048] Element <![CDATA[WO3]]> Co Ta Cr Ti content / % 88.7 / / / /

[0049] The leaching residue is cobalt hydroxide, which can be directly used in the preparation of lithium cobalt oxide cathode materials. Its XRD analysis has been provided. Figure 4 Other trace elements can be used as doping and modifying elements for lithium cobalt oxide to improve the electrochemical performance of the cathode material. The process includes the following steps: using filter residue as a precursor, a certain proportion of lithium source is weighed, and the lithium source and filter residue are mixed evenly and sintered at high temperature to obtain the lithium cobalt oxide cathode material. Lithium source and the leached filter residue Co(OH)2 are weighed according to a molar ratio of Li:TM = 1.05:1, mixed evenly, and sintered at high temperature in stages. First, calcination is carried out at 480℃ for 5 hours in a pure oxygen atmosphere, then the temperature is raised to 920℃ for 12 hours, and finally, natural cooling is performed to obtain the cathode material LiCoO2. The cathode material made using this embodiment is used to assemble a button battery, and electrochemical performance tests are conducted. At 25℃, at 0.1C (1C = 200 mA / g) within a voltage range of 3–4.3V, the discharge specific capacity at 1C is 161.3 mAh / g, and the capacity retention rate reaches 89.6% after 100 cycles.

[0050] Example 2

[0051] Prepare 15g of tungsten carbide (containing cobalt) waste and calcine it in an oxygen tube furnace at 850℃ for 5 hours. Then grind it into fine powder using a ball mill. Add 20% sodium hydroxide solution to the fine powder and leach it at 90℃ for about 3 hours with a solid-liquid ratio of 1:3. After leaching, separate the cobalt hydroxide slag and sodium tungstate solution.

[0052] The organic phase for extraction was prepared using N235:2-octanol:kerosene in a ratio of 10:10:80. First, acidify 50 ml of N235 with 25 ml of 15% dilute sulfuric acid. Then, prepare 500 ml of organic phase using 50 ml of2-octanol and 400 ml of kerosene. Acidify the filtrate to pH 4.5 with dilute sulfuric acid. Take 100 mL of the filtrate and 200 mL of the organic phase in a separatory funnel, shake for 2-3 minutes, and let stand. After separation, remove the aqueous phase and retain it. The extracted negative tungsten organic phase was back-extracted with 100 ml of 15% ammonia solution. After thorough mixing, allow it to separate into layers. Repeat the ammonia back-extraction process twice to obtain an ammonium tungstate solution. After removing the oil from the solution, evaporate and crystallize at 90°C to obtain ammonium paratungstate crystals.

[0053] Example 3

[0054] Prepare 30g of tungsten carbide (containing cobalt) waste and calcine it in an oxygen tube furnace at 900℃ for 8 hours. Then grind it into fine powder using a ball mill. Add 35% sodium hydroxide solution to the fine powder and leach it at 90℃ for about 4 hours with a solid-liquid ratio of 1:5. After leaching, separate the cobalt hydroxide slag and sodium tungstate solution.

[0055] The organic phase for extraction was prepared using N235:2-octanol:kerosene in a ratio of 9:11:80. First, 45 ml of N235 was acidified with 30% dilute sulfuric acid. Then, 55 ml of2-octanol and 400 ml of kerosene were used to prepare 500 ml of the organic phase. The filtrate was acidified to pH 3.0 with dilute sulfuric acid. 50 mL of the filtrate and 150 mL of the organic phase were placed in a separatory funnel, shaken for 5-8 minutes, and allowed to stand. After separation, the aqueous phase was collected and retained. The negative tungsten organic phase was back-extracted with 300 ml of 28% ammonia solution. After thorough mixing, the phase was allowed to separate. The aqueous phase was collected, and the ammonia back-extraction process was repeated twice to obtain an ammonium tungstate solution. After removing the oil from the solution, it was evaporated at 75°C to crystallize, yielding ammonium paratungstate crystals.

[0056] Example 4

[0057] Prepare 20g of tungsten carbide (containing cobalt) waste and calcine it in an oxygen tube furnace at 1100℃ for 10 hours. Then grind it into fine powder using a ball mill. Add 30% sodium hydroxide solution to the fine powder and leach it at 85℃ for about 3 hours with a solid-liquid ratio of 1:6. After leaching, separate the cobalt hydroxide slag and sodium tungstate solution.

[0058] The organic phase for extraction was prepared using N235:2-octanol:kerosene in a ratio of 10:15:75. First, 50 mL of N235 was acidified with 20% dilute sulfuric acid. Then, 75 mL of2-octanol and 375 mL of kerosene were used to prepare 500 mL of the organic phase. The filtrate was acidified to pH 1.5 with dilute sulfuric acid. 50 mL of the filtrate and 50 mL of the organic phase were placed in a separatory funnel, shaken for 5 minutes, and allowed to stand. After separation, the aqueous phase was collected and retained. The negative tungsten organic phase was back-extracted with 150 mL of 20% ammonia solution. After thorough mixing, the phase was allowed to separate. The aqueous phase was collected, and the ammonia back-extraction process was repeated twice to obtain an ammonium tungstate solution. After removing the oil from the solution, it was evaporated at 95°C to crystallize, yielding ammonium paratungstate crystals.

[0059] Table 1 W extraction rate under different extractants

[0060] Example Conditions (N235: 2-octanol: kerosene) Raffinate concentration (W) / g / mL Organic phase extraction rate / % 1 9：11：80 2.8 99.99 2 10：10：80 1.3 99.99 3 9：11：80 2.2 99.99 4 10：15：75 0.9 99.99

[0061] Example 5

[0062] Prepare 50g of tungsten carbide (containing cobalt) waste and calcine it in an oxygen tube furnace at 1100℃ for 6 hours. Grind it into fine powder using a ball mill, controlling the particle size to be approximately 20μm. Add 35% sodium hydroxide solution to the fine powder and leach it at 90℃ for about 6 hours, with a solid-liquid ratio of 1:4. After leaching, separate the cobalt hydroxide slag and sodium tungstate solution.

[0063] The leached sodium tungstate solution was concentrated to approximately 65 g / L. 100 ml of the concentrated sodium tungstate solution was mixed thoroughly with 10 ml of 98% concentrated sulfuric acid to allow the reaction to occur. Tungstic acid, which is poorly soluble in water, precipitated. The reaction solution was filtered to separate the solid and liquid phases. The tungstic acid precipitate was washed with 3% ammonia water and then dried at 80°C for 6 hours. The dried tungstic acid was added to 28% concentrated ammonia water with continuous stirring, maintaining the temperature at approximately 60°C and the ammonia concentration in the solution at 25-40 g / L. Once the tungstic acid was completely dissolved in the ammonia water, a mixed solution of ammonium tungstate and ammonia water was obtained. This solution was then evaporated and crystallized at 90°C to obtain ammonium paratungstate crystals. The tungsten recovery rate during the acid precipitation process was 86.74%. The results of the analysis of the leaching solution and the ammonium tungstate mixed solution are shown in the table below.

[0064] Example 5 Leachate Composition Table

[0065] Element W Co Ta Cr Ti Content / ppm 41835 0.2 / 0.03 0.9

[0066] Example 5: Composition table of ammonium tungstate mixed solution

[0067] Element W Co Ta Cr Ti Content / ppm 13389 / / / /

[0068] Comparative Example 1

[0069] Prepare 20g of tungsten carbide (containing cobalt) waste, mix it with sodium hydroxide powder, and calcine it in an oxygen tube furnace at 1000℃ for 6 hours. Then, put the calcined material into deionized water and heat it at 90℃ for about 5 hours with a solid-liquid ratio of 1:4. After leaching, separate the leaching residue and sodium tungstate solution.

[0070] The results show that the leaching effects of tungsten and cobalt are significantly worse than those of the two elements in the examples. Due to the high hardness of tungsten carbide, its texture gradually changes during sintering but it is still not easily broken (e.g., Figure 2 In the comparative example, the sodium hydroxide powder added did not easily penetrate into the particles for further reaction, and the tungsten carbide particles were large and uneven. During sintering, the uneven particle size prevented them from fully reacting with the sodium hydroxide powder. Therefore, the leaching effect in the comparative example was very poor.

[0071] Table 2 W / Co leaching rates under different leaching conditions

[0072] Example Conditions (NaOH concentration, leaching temperature, solid-liquid ratio, leaching time) W / ％ Co / % Example 1 25％，80，1:4，4 99.54 99.99 Example 2 20％，90，1:3，3 98.83 99.98 Example 3 35％，90，1:5，4 98.75 99.98 Example 4 30％，85，1:6，3 98.76 99.95 Comparative Example 1 90，1:4，5 58.37 /

[0073] The method provided by this invention can obtain high-purity cobalt and tungsten through a simple process of oxidation calcination, sodium hydroxide leaching, and extraction crystallization. Compared with conventional roasting-alkali leaching, electrolysis, zinc smelting, and nitrate smelting methods, this invention does not involve a complex sintering process, has a simple recycling process, and does not require expensive, high-requirement processing equipment, making it easy to industrialize. At the same time, the required raw and auxiliary materials are simple and readily available, resulting in low production costs, less wastewater, no waste gas generation, and minimal environmental pollution. Furthermore, the product has high purity and high recycling value, offering excellent economic benefits. Unlike traditional tungsten ore alkaline leaching, the method provided by this invention can leach at a low temperature and normal pressure at 100℃, with a leaching efficiency of over 98%. This method is suitable for the efficient recovery of all transition metal modified tungsten carbide waste, including cemented carbide cutting tools and drill bits.

[0074] The present invention has been described in detail above with reference to specific embodiments and exemplary examples; however, these descriptions should not be construed as limiting the present invention. Those skilled in the art will understand that various equivalent substitutions, modifications, or improvements can be made to the technical solutions and embodiments of the present invention without departing from the spirit and scope of the invention, and all such modifications and improvements fall within the scope of the present invention. The scope of protection of the present invention is defined by the appended claims.

Claims

1. A simple and efficient method for recycling tungsten carbide waste material, characterized in that: Includes the following steps, (1) Prepare tungsten carbide waste and calcine it at different temperatures to destroy its structure; (2) Grind the completely calcined tungsten waste into fine powder; (3) Add alkaline solution to the calcined and ground fine powder, leach it, and then separate the solid and liquid to obtain sodium tungstate filtrate and filter residue; The calcination temperature in step (1) is 600-1500℃, the calcination time is 2-50h, and the calcination atmosphere is an oxygen-containing atmosphere; The alkaline solution in step (3) is a sodium hydroxide solution with a concentration of 20-46%, a solid-liquid ratio of 1:1-10, a leaching temperature of 60-100℃, and a leaching time of 1-12h. The filter residue in step (3) contains cobalt hydroxide, and the process also includes mixing and sintering a lithium source with the filter residue to prepare a lithium cobalt oxide cathode material; The molar ratio of the lithium source to the metal in the filter residue is Li:TM = (1.03-1.25):

1.

2. A simple and efficient process for recycling cemented tungsten carbide waste material according to claim 1, characterized in that: Ammonium paratungstate is prepared by extraction and crystallization with sodium tungstate solution after leaching, specifically including the following steps: S1 Extraction Organic Phase Preparation: First, acidify N235 with dilute sulfuric acid, then prepare the organic phase using 2-octanol and kerosene. S2 extraction: Acidify the sodium tungstate filtrate, mix the acidified solution thoroughly with the organic phase, let it stand, and after standing and separating the layers, separate the aqueous phase and retain the negative tungsten organic phase; S3 back-extraction: The negative tungsten organic phase is back-extracted with ammonia water, so that the tungsten anions in the organic phase enter the aqueous phase in the form of ammonium tungstate, thereby achieving the purpose of converting sodium tungstate into ammonium tungstate and separating it from the organic phase; S4 Evaporation and Crystallization: The ammonium tungstate solution after back-extraction is de-oiled and then evaporated and crystallized to obtain ammonium paratungstate.

3. A simple and efficient method of recycling cemented tungsten carbide waste material according to claim 2, characterized in that: The concentration of dilute sulfuric acid in step S1 is 5%-35%, and the proportions of N235, 2-octanol, and kerosene are 5-10%, 10-20%, and 70-85%, respectively, with a total of 100%.

4. The method of simple and efficient recycling of cemented tungsten carbide waste material according to claim 2, characterized in that: The solution in step S2 is acidified to pH 1.0-5.0, and the ratio of the aqueous phase to the organic phase is 1:1-6.

5. The simple and efficient method for recycling cemented carbide tungsten carbide waste according to claim 2, characterized in that: The ammonia concentration in step S3 is 10%-28%, and the ratio of ammonia to organic matter is 1-6:

1. The evaporation and crystallization temperature in step S4 is 70-150℃.

6. The method of simple and efficient recycling of cemented tungsten carbide waste material according to claim 1, characterized in that: The lithium source is one or more of lithium hydroxide, lithium carbonate, and lithium nitrate.

7. The method of simple and efficient recycling of cemented tungsten carbide waste material according to claim 1, characterized in that: The sintering process involves a first-stage calcination temperature of 450-600℃ and a holding time of 4-8 hours; and a second-stage calcination temperature of 850-1000℃ and a holding time of 9-16 hours. The calcination atmosphere in these steps is one of air, oxygen-enriched air, or pure oxygen.

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